Process to produce simonkolleite, zinc oxide and zinc hydroxide

ABSTRACT

A hydrometallurgical process utilizing an atmospheric calcium chloride leach to selectively recover from various metal feed stocks (consisting of elemental metals, metal oxides, metal ferrite, metal hydroxide, metal carbonates, metal sulfate/sulfur compounds, and their hydrates, specifically including but not limited to EAF Dust K061) zinc, lead, cadmium, silver, copper and other valuable metals to the exclusion of iron, magnesium, halogen salts and other unwanted elements. The process solves the problem of iron and magnesium leach solution contamination because iron is unexpectedly converted to magnetite. The heavy metals are cemented out of solution using zinc or other selected dust at a pH of 6 or greater under unique and unexpected conditions, which do not require acid. Simonkolleite/zinc- oxychloride/zinc-hydroxide is produced from the purified zinc chloride complex pregnant leach solution and is converted directly to high purity active rubber grade 99+% zinc oxide having small particle size and high surface area. The products are metal concentrates suitable for: metal refiner/processors, production of elemental metal, or other conversion processes. The process removes Arsenic and Fluorides in the feed material. The process also solves the problem of chloride contamination in the zinc oxide and prevents heavy metal contaminants in the hydrometallurgically produced zinc oxide derived from feed stocks containing chlorides or when chlorides are used to leach the metal bearing feed stocks. In one embodiment, calcium and/or magnesium compounds are added to the iron bearing waste to increase the recovery of zinc and other non-ferrous metals and to produce an iron bearing flux. The process is environmentally friendly and fully recycles all streams.

CROSS-REFERENCE TO RELATED APPLICATION

This application is a continuation of U.S. application Ser. No.09/478,625, filed Jan. 5, 2000, now U.S. Pat. No. 6,770,249, which inturn claims the benefit of U.S. Provisional Application No. 60/156,163,filed Sep. 27, 1999.

FIELD OF THE INVENTION

The present invention relates to a method and apparatus for theeconomical processing of metallurgical dust, such as electric arcfurnace (hereinafter EAF) dust, also known as Environmental ProtectionAgency (hereinafter EPA) hazardous waste No. K061, into marketablechemicals and products. EAF dust, which is often treated as hazardouswaste, can now be a potential raw material for the recovery of the metalvalues contained therein as metals, inorganic chemicals, and as apotential raw material source.

BACKGROUND OF THE INVENTION

The manufacture of steel using an electric arc furnace is a highlyadvantageous process in the modern steel industry. A drawback in the EAFmanufacturing of steel is the production of EAF dust waste by-product,which is an EPA listed hazardous waste (K061). During melting andrefining of steel in electric arc furnaces, large amounts of galvanizedscrap material can be fed into the EAF process. Inherent in the processfor making steel by the use of electric steel furnaces with submergedarcs is the liberation of zinc, iron, and other metal values as EAF dustin the off-gas leaving the furnace. To protect the atmosphere, theseparticles are removed in baghouses, cyclones, scrubbers, and othersimilar devices. Due to the high proportion of zinc in the dust, it isespecially desirable as a source of zinc values for sale. The dustconsists of fine particles of iron oxides, ferrites, calcium oxide, andsilica, metal chloride and oxide particles of nonferrous metals such aszinc, lead, cadmium, and silver, which vaporize at the high temperaturesof the molten steel bath, and which are recovered in the dust.

The feeding of fine particles, containing unwanted metals, back to thearc furnace is not economically viable. As the amount of recycled dustincreases, the energy requirement for reducing and melting the iron andother metals from the dust increases, causing melt chemistry problemsand decreasing furnace refractory life. Traditionally, this dust hasbeen considered a waste material and has been disposed of in landfills.Recovery techniques used previously generate waste by-products orproducts not suitable for their intended claimed use and do not offersufficiently high recovery yield of the metals and products in the EAFdust.

The rapid growth of the EAF steel process has made EAF dust one of thefastest growing and largest environmental problems worldwide. Thelandfill disposal method is becoming more expensive because ofincreasingly stringent Environmental Protection Agency (EPA)regulations. The chemical nature of these dust particles are such thatthey classify as listed hazardous waste, based on the toxicity testprescribed by the United States Environmental Protection Agency. Thetoxicity concern is related to the presence of lead, cadmium, andchromium.

At present, there are approximately 925,000 tons of this hazardous wastegenerated annually in the United States and an additional 3,000,000 tonsgenerated annually in the rest of the world. EAF produced steelcomprises forty-five percent (45%) of the total U.S. steel production.It is expected to become the major source of steel produced in the U.S.within the next few decades. At present, approximately one-half of theU.S. production of EAF dust is being land filled.

It is a primary purpose of this invention to recover metal values fromthis steel-making flue dust, and particularly to recover zinc oxidesuitable for rubber compounding, and additionally to provide a means forthe separation and recovery of other materials in the dust, with minimalenvironmental impact.

There is also a similar but lower concentration zinc contaminated dustwhich is derived from the other major process for steel manufacturing,the basic oxygen furnace or basic oxygen process (hereinafter BOF).Because the levels of toxic metals such as cadmium, lead and zinc arelower than current toxicity cutoff levels, BOF dust is not currentlyclassified by the EPA as hazardous. However, BOF dust may be classifiedas hazardous in the future and its non-iron contaminants, like zinc,make it difficult to utilize in current steel manufacturing, resultingin substantial worldwide stock piles of BOF dust.

With the resource of EAF and BOF dust readily available, and based onenvironmental need, this process was developed to economically recovernonferrous metals such as zinc, lead, and cadmium from these steel plantdusts. The iron oxide, depleted of these metals, can be recycled back tothe steel furnace. Since the tonnage of this raw material issubstantial, it represents an important source of zinc, lead and ironmetals.

There are also waste dusts and metal sludges available from zinc andcopper recovery and extractions processes and other metal processeswhich also represent valuable sources of non-ferrous metals.

Accordingly, there is a need to develop a hydrometallurgical chloridebased process to economically recover valuable metals from metal bearingwaste and ores which will solve the problem of unwanted ioncontamination of the pregnant leach solution, thus insuring thatfinished products have extremely low levels of contamination of unwantedions. Chloride leach processes offer apparent advantages over sulfurbased chemistry, such as avoiding roasting, sulfuric acid regeneration,and unwanted waste by-products. Even though these processes offer theopportunity to be environmentally friendly, they have not had mucheconomic success. The need to find a chloride based chemistry process isespecially crucial for metal bearing feed stocks that contain chloride.Previously developed hydrometallurgical chloride based leachingprocesses have required the addition of costly metal chloride saltadditives (components like ferric chloride or cupric chloride) orhydrochloric acid, which cause unwanted ions in the pregnant leachsolution that have to be removed prior to subsequent metal extractionrecovery steps. Other previously developed chloride processes containammonium, with resulting safety concerns and the potential for unwantedammonium compounds. These currently available processes utilize hightemperatures, high pressures, and/or highly acidic conditions.

The prior art, as described in U.S. Pat. No. 1,863,700, teaches thatzinc oxide/oxychlorides produced from simple precipitation with calciumhydroxide contain as little as 1% chlorides and 3-5% calcium. The patentfurther teaches a process where a zinc product can be produced with 1-2%calcium and 0.8% chlorides. Such a product thus contains 7%Simonkolleite or other oxychlorides based on the chloride stoichiometryand thus contains less than 92% pure zinc oxide, which is unacceptablefor rubber grade zinc oxide and other high purity zinc oxides as well aschloride intolerant zinc refining processes. In contrast, the process ofthe present invention converts Simonkolleite and other zinc oxychloridesto zinc oxide, meeting specifications and suitability for use as rubbergrade zinc oxide with more than 99% pure zinc oxide. The zinc oxide hasless than 1,800 ppm chlorides and usually less than 1000 ppm and has aparticle size of 0.05 micron to 0.5 micron. The process produces anactive zinc oxide with a surface area of 10-70 m²/gram, providing betterreactivity and economy compared to French processed zinc oxide. (See“Active Zinc Oxide—the Advantage” by Dr. Harry Rothmann and L.Bruggemann-Sprit und Chemische Fabrik, Germany Tire Technol. Int., p.118 (1997)). This large amount of surface area offers a more reactiveproduct allowing its use at lower levels than is currently practiced,with improved economics and in processes where French Process zinc oxidehas not been acceptable and/or viable.

In the present invention, a pyrolysis process is described where in aniron bearing material containing non-ferrous metals (specificallyincluding but not limited to EAF Dust-KO61) is reduced of its oxides andefficiently stripped of its non-ferrous metals in the presence of carbonand calcium; and a hydrometallurgical process is described utilizing anatmospheric calcium chloride leach to selectively recover from variousmetal feed stocks (consisting of elemental metals, metal oxides, metalferrites, metal hydroxides, metal carbonates, metal sulfate/sulfurcompounds, and their hydrates, specifically including but not limited toEAF Dust-KO61) the following components: zinc, lead, cadmium, silver,copper and other valuable metals to the exclusion of iron, magnesium,halogen salts and other unwanted elements. The pyrolysis process solvesthe problem of high levels of non-ferrous metals left in the iron richmaterial produced from KO-61 when not using calcium.

SUMMARY OF THE INVENTION

The process utilizes a chloride leach solution for leaching selectivemetals from metal bearing feed stocks, producing a pregnant chlorideleach solution suitable for selective metal recovery or other metalextraction which removes, separates, and recovers such metals in a costeffective manner with minimal amounts of unprocessed solids and sludgeremaining in the environment. The process is based on the addition ofwater, rather than the use of an acid, to control the chlorideconcentration. In addition, the process produces an intermediaryproduct, Simonkolleite, which is then processed to produce 99% plusgrade zinc oxide. The hydrometallurgical process utilizes an atmosphericcalcium chloride leach to selectively recover from metal bearing feedstock zinc, lead, cadmium, silver, copper and other valuable metals tothe exclusion of iron, magnesium, halogen salts and other unwantedelements. The pregnant leach solution is ideally suited to makehigh-grade zinc oxide and recover other heavy metals with minimalcontaminants utilizing cementation processes or selective ion extractionutilizing liquid-liquid extraction processes. It has the unexpectedbenefit of converting iron oxide and ferrites to magnetite.

The process has been specifically designed to fully recycle all processliquids while producing marketable solid products.

BRIEF DESCRIPTION OF THE DRAWINGS

The foregoing aspects and many of the attendant advantages of thisinvention will become more readily appreciated as the same become betterunderstood by reference to the following detailed description, whentaken in conjunction with the accompanying drawings, wherein:

FIG. 1 is a schematic flow diagram of the zinc oxide recovery processutilizing lime for zinc precipitation and hearth dust as the feed stock;

FIG. 2 is a schematic flow diagram of the zinc oxide recovery processutilizing caustic for zinc precipitation and hearth dust as the feedstock;

FIG. 3 is a schematic flow diagram of the zinc oxide recovery processutilizing lime for zinc precipitation and EAF dust as the feed stock;

FIG. 4 is a schematic flow diagram of the zinc oxide recovery processutilizing lime for zinc precipitation and EAF dust as the feed stock;

FIG. 5 is a schematic flow diagram of the zinc oxide recovery processutilizing lime for zinc precipitation and EAF dust as the feed stock;

FIG. 6 is detailed schematic flow diagram of the zinc oxide from zincoxychloride portion of the overall process; and

FIG. 7 shows a graph which compares the zinc oxide produced by thepresent inventive process with zinc oxide prepared by a standardprocess;

Tables 1-18 show data related to the present invention.

DETAILED DESCRIPTION OF THE PREFERRED EMBODIMENT

A general overview of the process will first be described.

Electric Arc Furnace: Scrap iron, iron ore, carbon, calcium oxide,magnesium and oxygen along with other feed stock, is fed to a furnace.Two electrodes (probes) are used to melt the metal. Dust is emitted withmetal vapor which turns to oxides, known as EAF Dust (KO-61). TheElectric Arc Furnace could also be a blast furnace producing any type ofsteel. The process may comprise any metal furnace or metal process whichmakes dust or sludges with recoverable metals.

Rotary Hearth Furnace: The K061 may contain iron lead, copper, cadmium,silver and zinc and is mixed with carbon, and may further be mixed withcalcium and magnesium rich stocks. This mixture is fed to the furnaceand its metals are reduced (stripped of their oxides) in an oxygenreducing atmosphere. The non-volatile metal makes direct reduced ironand calcium/magnesium briquettes and a Hearth Dust rich in non-ferrousmetals. Hearth dust includes zinc oxide, lead, copper, silver, chloridesand cadmium. The inventive process can be used with an input of KO-61,metal rich Dust or sludges or Hearth Dust. In other words, the hearthfurnace can be positioned at the beginning or at the end of the process.In addition, the process may use lime and/or caustic in the chemistry.In the process using caustic, the caustic is regenerated or can be soldas sodium/potassium chloride.

Leach Reactor: zinc, lead, copper and cadmium are fed in to the reactor,along with other heavy metals and mixed with a leach solution. The leachsolution is typically 53% calcium chloride and 47% water, but maycontain sodium and potassium chloride up to their respective saturationlevels (generally 1-7%). This is a straight salt leach with no acidadded. In this leach step, iron, magnesium, calcium and other relatedelements do not get absorbed into the leach.

Filter Leach Slurry: The dregs/filter cake from the leach filter includecalcium hydroxide, which has been separated from the valuable metals,and is shipped for sale or further processing back to the furnace. Thefiltrate liquid includes the valuable metals including specificallyzinc, lead, copper, silver and cadmium. There are very few contaminantsin this pregnant filtrate stream.

Cementation: This step separates the lead, copper, cadmium and silverfrom the zinc pregnant leach solution with zinc powder in thecementation step. This step uses zinc to displace the lead, copper,cadmium and other similar metals from a metal chloride complex. In priorart industry processes, this step is conducted in the range of pH 2-6.The inventive process, however, uses a water dilution instead of directpH adjustment. In other words, water is added to the solution to changethe chloride concentration which allows the reaction to occur at anunadjusted pH, typically at a pH of 6-8. This step produces elementalmetals (cement), which generally comprises lead, cadmium, silver andcopper. The cement is fed to a lead smelter or further refined andseparated by other known processes.

Zinc Precipitation Step: Water is added to the zinc pregnant solutionwhich produces a white precipitation called Simonkolleite. Approximately35% of the Simonkolleite precipitates by the addition of water. Lime issimultaneously or subsequently added to precipitate out the remaining65% of the Simonkolleite.

Chemistry: The unique chemistry of the inventive process is:2MO+3CaCl₂+2H₂O Ca⁺⁺+[MCl₃ ⁻]₂+Ca (OH)₂,  Equation 1wherein M=Zn, Pb, Cu and/or Cd, and wherein M in the chloride complex isdisplaced with Zn prior to the reaction of Equation 2.Ca⁺⁺[ZnCl₃ ⁻]₂+Ca(OH)₂ZnCl₂.4(Zn(OH₂).H₂O+ZnO,  Equation 2wherein the ZnCl₂.4(ZnOH₂).H₂O is Simonkolleite, containing equilibriumamounts of ZnO and Zn(OH)₂

In cold temperatures, the production of Simonkolleite is increasedwhereas at hotter temperatures, more ZnO is produced. Accordingly, theabove reaction preferably takes place at ambient temperatures toincrease the production of Simonkolleite and to reduce the production ofZnO. (Note: There is no temperature at which only zinc oxide can beproduced given the elevated levels of chlorides in solution.) Thedesirability of using cold temperatures to produce Simonkolleite iscounter intuitive because the production of Simonkolleite decreases theamount of zinc oxide produced in this step. Production of theintermediary Simonkolleite, however, is then used to produce zinc oxideof high purity and with less environmental waste than prior artprocesses, compared to a process wherein zinc oxide is produced directlywithout the intermediary.

Simonkolleite is filtered from the solution. The filtrate is evaporatedso that NaCl and KCl crystallize out and the solution is suitable to berecycled for use as leach solution.

The Simonkolleite is then reslurried with fresh water or recycled waterand reacted with lime or caustic at elevated temperatures and pressuresto produce the zinc oxide.

We now refer to the figures for a more detailed description.

FIG. 1 shows schematically a process 10 of a preferred embodiment of thepresent invention. Beginning at the front end of the process in theFurnace section, an electric arc furnace (EAF) 12 outputs EAF dust in astream 9900. This dust is classified by the EPA as K061, a listedhazardous waste. The metal bearing feed stock typically is chosen fromthe group of materials including: metal furnace dusts, smelting dusts,metal refining dust, metal bearing waste sludges, mill tailings, oreswhich contain metal oxides, metal hydroxides, metal ferrites, metaloxide and their compounds, metal sulfates/sulfites/sulfur compounds,carbonates, and metal bearing materials containing chlorides orfluorides. The metal bearing feed stock may also comprise Electric ArcFurnace Flue dust K061 derived from the off gasses of the Electric ArcFurnace processing of scrap steel where said dust contains approximately10-40% zinc or the Electric Arc Furnace dust can contain zinc ferrite.The feed stock can also comprise, but is not limited to, zinc and metaloxide dust (zinc concentrate) recovered frompyrolysis/furnace/kiln/roaster operations to roast or reduce (metallize)metal bearing ore or waste metal bearing materials such as EAF Dust whenmaking direct reduced iron, sludges from metal plating baths, sludgesfrom electrowinning tank operations, smelter furnace dust from zinc,copper or other heavy metal, recover/processing, iron and steelproduction including blast furnace or BOP or BOF flue dust containing3-20% zinc and heavy metal content, Jarosite from mining extractionoperations, waste water treatment metal bearing sludges or cakes, zincferrite tailings, EAF dust containing 10-60% or more zinc, zinc/coppersmelter dust, iron ore containing greater than 5% zinc and heavy metals,blast furnace/BOP dust containing 5-35% or more zinc, zinc smelter dust,copper smelter dust, metal bearing ores containing iron or sulfates,metal concentrates containing sulfur compounds, computer and electroniccomponent recycling dusts and sludges, waste by-product streamscontaining oxides or hydroxides, metal bearing materials containingchlorides, copperhead dross, cyanide bearing plating waste, platingwaste, arsenic bearing ores and wastes, gold ores and their by-productswith sulfur, copper electrolysis anode sludge, ores of all metals,incinerator fly ash, boiler fly ash, spent metal catalyst, metallicslags/dross, uranium ores, gold bearing ore, silver bearing ore andsludges, feed stocks or solutions containing fluoride, feed stocks orsolution containing arsenic, galvanized scrap metal, zinc bearing scrap,copper bearing scrap, scrap electronic components, circuit boardrecycling, and/or KO-88 aluminum pot liner. The feed stock of stream9900, typically comprises EAF Dust having a composition of approximately20% zinc, 6% lead, 40% iron, and various other miscellaneous components.The feed stock may be shipped to the processing plant from the steelmill by truck or rail.

Stream 9900 is fed to a hearth furnace 14, along with stream 1600 from aleach filter 16. The hearth furnace turns the oxides into metals by theaddition of coke (carbon). Specifically, K061 plus coke or other carboncontent, in reducing conditions, causes the oxides to depart from themetals leaving metallic iron. The zinc, lead, copper and cadmiumvaporize off as a metal. In the preferred embodiment of the processcalcium hydroxide or oxide and magnesium oxide or hydroxide is added tothe iron bearing material prior to metallization/reduction to enhancethe non-ferrous metal recovery yields and provide a more pure ironcalcium rich material for iron and steel production.

Stream 1900 comprises 80-99% metallized iron briquettes fed back to theEAF. The stream also includes calcium oxide (reduced from calciumhydroxide) and magnesium oxides. Stream 1900 is called “fluxed sinter”,also referred to as direct reduced iron (DRI).

Hearth dust is fed from hearth furnace 14 in stream 9950 to hearth duststorage bin 18. The hearth dust is a gray/white dust that is producedwhen the zinc metal vapor and other metals are re-oxidized in the top ofhearth furnace 14. The dust comprises 30 to 80% zinc, and typically 50to 60% zinc. The other components comprise lead, copper, cadmium,silver, other recoverable metals, chlorides and trace contaminants suchas iron and manganese. Stream 9950 is fed to storage bin 18. Stream 100is fed from storage bin 18 as needed such that stream 100 has the samechemical make-up as stream 9950.

Hearth dust is fed from storage bin 18 in stream 100 to a leach reactor20. A stream 200 is also fed to reactor 20 and comprises a calciumchloride leach solution with a specific gravity of approximately 1.49.The solution of stream 200 may also comprise sodium chlorides andpotassium chlorides in amounts from a trace amount up to theirsaturating level of approximately 1-8% of the solution. The calciumchloride typically is recycled from another portion of the process. Inreactor 20, the zinc oxide in the hearth dust combines with the calciumchloride to produce a zinc chloride complex and a calcium hydroxideprecipitate. Other metal oxides in the dust will also combine with thecalcium chloride to produce a corresponding metal chloride complex. Theprocess simultaneously follows three separate equations:3CaCl₂+2MO+2(H₂O)[Ca⁺⁺+2MCl₃ ⁻]+2Ca(OH)₂  Equation 3CaCl₂MCl₂[Ca⁺⁺⁺MCl₄ ⁻]  Equation 4CaCl₂+MO+H₂O MCl₂+Ca(OH)₂,  Equation 5wherein M=Zn, Cd, Cu, Ag, Sn, Ni and/or Pb, and wherein the metalchloride complex is MCl₃ ⁻.

Approximately 95%+ of the feed solution undergoes the reaction ofEquation 3 so that approximately 5% undergoes the reactions of Equations4 and 5 combined. Accordingly, stream 300 comprises 95% of the metal aschloride complex and calcium hydroxide solids and dregs (the unleachedmaterials) including magnesium, manganese and iron. The specific gravityof the materials fed to filter 16 in stream 300 preferably isapproximately 1.49.

Stream 300 is fed to a leach filter 16, which typically comprises apress or other liquid-solid separation device. During the batch processdisclosed herein, this filter processing step comprises three distinctsteps. In a continuous process, three separate filters may be used.

In the first filtering step, filter 16 presses the solution to removethe liquid from the solids of stream 300. This liquid is removed asstream 600. In the second filtering step, the solids remaining in thefilter are washed with stream 210, comprising calcium chloride solutionfrom tank 22 or its dilute form, and pressed to recover additionalmetals from the filter cake. The liquid is removed as stream 650. Thisstep increases the recovery of metals from the filter. In the thirdfiltering step, the solids remaining in the filter are washed withstream 500, comprising hot water, to further improve the recovery ofmetals into the liquid solution from the filter cake. These three liquidstreams are mixed together as stream 675 and comprise the metal chloridecomplexes, MCl₃ ⁻. These liquid streams should be held above thecrystallization temperature of lead chloride, and typically are held ata temperature of approximately 90° C. The solids, i.e., the filter cake,are sent via stream 1600 to hearth furnace 14. The solids compriseapproximately 80% calcium hydroxide and approximately 20% magnesium andother dregs. The solids may be dried before being returned to thefurnace to be calcined. Stream 1600 may be sold to outside vendorsrather than being recycled back to the hearth furnace as a calciumhydroxide or calcium oxide rich stream.

Leach filter 16 feeds streams 625, 650 and 600 to stream 675, which inturn is fed to a heavy metal cementation tank 24. Stream 675 may bepassed through an iron removal process 21 (FIG. 3) before being passedto cementation tank 24 to remove any trace amounts of iron in thestream. Stream 675 typically comprises a green liquid containing zinc,lead, copper, cadmium and silver in the chloride complex form.

In the preferred embodiment, the chloride leach solution of stream 200is calcium chloride, and may also contain sodium chloride and potassiumchloride or other chloride salts from groups 1 or 2 on the periodictable, below or at their solubility level in calcium chloride at thechosen concentration. The leach solution chloride ion concentrationtypically is greater than 10 molar and less than the saturation point atthe chosen operating temperature for calcium chloride. The calciumconcentration preferably is at least 16% and less than the saturationpoint for calcium chloride. The leach solution typically is fully liquidat the selected temperature and below the saturation point. The leach ofreactor 20 preferably occurs at atmospheric pressure, but may beconducted at higher pressure conditions. The feed stock particles shouldbe sufficiently small to allow good reaction rates. The reaction may bewell agitated to decrease reaction times. During the reaction, the pulpdensity typically is 1% or greater. The temperature preferably is above70° C. and below the boiling point of the solution at the selectedoperating pressure and at a temperature at which all metal chloridespresent are soluble. The calcium concentration in the leach solution ofreactor 20 may be, but is not required to be, maintained/reestablishedby the addition of HCl to the leach cake containing calcium hydroxidesor other calcium complexes precipitated during the reaction. When soelected, the pH should be maintained above 3.5 and below 9.0. Thecalcium content may be reduced by a reaction with sulfuric acid on theleach solution and/or on the recycled leach solution feed to producegypsum. The reaction typically occurs in the pH range of between 3.5 to9.5 without the addition of acid. The primary metals targeted forrecovery are zinc, lead, cadmium, silver, copper, nickel, tin, and othermetals in the electrochemical series above zinc, including GroupClassifications 9, 10, 11, 12, 13, 14, 15. The process produces apregnant leach solution from which one can produce elemental metals,metal oxides or metal hydroxides directly, extract selective ionsutilizing liquid-liquid or resin exchange, or produce metal chloridesfor electrowinning elemental metal. The leach can be carried out as asingle stage or a multi-stage process, a continuous process, and may beco-current or countercurrent.

The hydrometallurgical process rejects substantially all the iron, Mg,Cr, calcium hydroxide, fluoride, Ca/Mg/Ba, carbonates and hydroxides,sulfates, (and other group 1-8 elements), carbon, Si, As and P can berejected (not leached) based on the concentration of chloride ions, thetemperature, and other operating conditions. The process solublizesapproximately only 2-100 ppm iron in the leach solution. Unexpectedly,the process converts Fe₂O₃ and other iron compounds to magnetite Fe₃O₄.The magnetite can be used as an iron ore substitute, feed for directreduced iron production, or may be used for coal ore floatation and acementation additive. The undissolved solids of the leach process areseparated (i.e. filtered, centrifuged, etc.) out of the pregnant leachsolution and washed first with leach solution and then with water toobtain an enhanced metal yield recovery and purified cake. The solidstypically are washed with hot chloride leach solution to removeentrained metal chloride complexes. The wash temperature typically isgreater than the crystal formation point for the chloride salts present.The leach filter cake is washed with a series of hot leach solutions ofreduced (or increased) chloride concentrations to recover metals thatdid not dissolve at the chosen operating concentration conditions. Thesolids are then washed with hot water above 80° C. and at a temperaturegreater than the crystal formation point for the chloride salts present,to remove the remaining entrained leach wash solution. The solids arewashed with fresh/recycled hot water to remove soluble chloride salts.The leach typically recovers 70-97% of all heavy metals from the feedmaterial. The leach process does not use acid unless it is required aspart of a recycle loop to maintain the chloride balance. The processdoes not require oxygen unless the rejected metals in the leach cake arepreferred in the fully oxidized form. The resultant pregnant leachsolution of the process contains low levels of Mg and Mn, whichminimizes contamination of products to be produced from the pregnantleach solution. Leach conditions of the process make magnetite Fe₃O₄when the feed material contains or produces iron oxide Fe₂O₃, ferritesor iron hydroxides. The inventive process removes Fluorides in the formof CaF₂. Fluorides arc made insoluble during the process and thereforeare removed from solution. Fluorides are converted to insoluble calciumfluoride which is environmentally stable. The salt leach willprecipitate NaCl and KCl to maintain a system balance. Arsenic reportsto the leach cake as an iron complex making the process environmentallysafe. During the process if cyanide is present it is complexed with ironto produce Prussian blue or decomposes to form carbon dioxide, makingthe process environmentally safe. By adding O₂ in a subsequent stage tothe leach process, the process selectively precipitates MnO₂ so that O₂is not added in an earlier step. Based on the belief, during theprocess, a platinum group metal, if present, selected from the groupconsisting of platinum, palladium, rhodium and ruthenium in an ore orother metal bearing feed, can also be recovered.

Still referring to FIG. 1, zinc powder in stream 9400, water includingsome chlorides such as zinc chlorides in stream 2500, water in stream570 and the metal chloride complexes in stream 675 are all fed to aheavy metal cementation tank 24. The water added to cementation tank 24in streams 2500 and 570 is added in precise amounts in order to controlthe chloride concentration in the tank. The water preferably is added tomaintain the specific gravity of the contents of the tank within a rangeof 1.41 to 1.49. By maintaining the specific gravity of the solutionwithin the range of 1.41 to 1.49, the zinc powder added to the solutionwill result in the zinc displacing each of the metals in the metalchloride complexes. A two stage counter current flow is preferred toreduce zinc dust addition. The reaction is as follows:MCl₃ ⁻+Zn ZnCl₃ ⁻+M,  Equation 6wherein M=Cd, Cu, Pb, Sn, Ni, Ag, or any other metals in chloridecomplex form, and wherein the metals precipitate out after beingdisplaced by the zinc.

The sequence of displacement by the zinc of particular metals can becontrolled by controlling the specific gravity of the solution. Inparticular, at a specific gravity of 1.49, the zinc selectivelydisplaces silver but does not displace any other metals.Correspondingly, other metals are believed to be displaced at otherparticular specific gravities. Accordingly, the addition of preciseamounts of water to the solution in sequential steps is believed toresult in the selective and sequential displacement by zinc ofparticular metals as desired.

Heavy metal cementation tank 24 feeds the displaced metals in stream2000 to outside sales or to a metals recovery loop. The displaced metalsare separated within tank 24 from the zinc chloride complex byfiltration and/or other liquid-solid separation methods. In a continuousprocess, the filtration and/or separation may take place in a series ofseparation filtration or gravity separation tanks. Typically the metalsolids are washed with water to remove the chloride salt solution. Afterwashing and drying, if being sold to outside sales, the solids arebriquetted in a press under high pressure and sold. If the displacedmetals are further processed instead of being sold directly, stream 2000is fed to cadmium leach tank 42, as will be further described below.

Referring to the zinc cementation process, the process comprisesextracting and recovering metals from a high chloride salt solutioncontaining metal ions by reducing the chloride content to specificlevels by the addition of water and then adding elemental metal to causemetals above said metal on the electrochemical replacement series tocement out of solution. Elemental zinc powder/dust and other selectivemetal dust is added to the pregnant leach solution, which is high inchloride content including but not limited to lead, copper, cadmium,silver, etc., and ions/complexes containing high chloride concentration,to unexpectedly and selectively, at a pH of 5 or greater, cement outmetals by a substitution reaction. The reaction form is: metallic dust(such as zinc)+metal chloride salt or metal chloride-complex=metal+zinc(or other metal) chloride or zinc chloride-complex ions. By varying thechloride concentration, instead of varying the pH, through the additionor removal of water, the expected and unwanted formation of metal oxidesand hydroxides are avoided and no acid or base addition is necessary tocontrol the pH to prevent precipitation of such metal oxides orhydroxides. The high chloride salt solution typically is calciumchloride. The temperature typically is above the crystallization pointof the solution and below the boiling point. The high chloride saltsolution may also contain sodium and/or potassium chloride up to theirrespective solubility limits in the calcium chloride solution at thechosen operating temperature and pressure. The solution typicallycontains more than 30% chlorides. The elemental metal used forcementation should be chosen from those metals below the metal to becemented on the electrochemical replacement series. Elemental zinc dustor zinc powder is used to cement lead, copper, cadmium, iron (only traceamounts are present, such as 6 ppm), mercury, silver and gold.Alternative cementation steps may include elemental copper dust orcopper powder to cement mercury, silver and gold prior to using zinc ina final cementation step. The metals being cemented out include, but arenot limited to lead, cadmium, copper, silver, tin, nickel, mercury,platinum and gold. The unexpected result of the process is that acontrolled lowering of the specific gravity or the chlorideconcentration allows the cementation of metals with elemental metal dustto occur at a pH above 4 without the co-precipitation of metal oxide,metal oxychloride and metal hydroxides, as one would normally expectunder these conditions. The temperature typically is held above thecrystallization point for all chloride salts present. The specificgravity is controlled between 1.40 and 1.49 to control chlorideconcentration by the addition of water. The elimination of oxygenfurther prevents the formation of zinc oxide, zinc oxychloride or zinchydroxide at temperatures below 80° C. The process results in less than1 ppm of each metal that is cemented out, being contained in thesolution after processing. The chloride concentration can alternativelybe controlled by selective chloride ion reduction, reaction chemistry orliquid-liquid extraction using prior art processes, instead of theaddition of water as discussed above. The process may be conducted inone stage, multiple stages, or continuously in a counter flow orco-current flow process. The metal dust used in the process preferablyis zinc. The amount of zinc dust added is between 1.1 and 2.0stoichiometric for the metal to be removed. The reaction is performed atatmospheric pressures and below the natural boiling point of thesolution. The metal sponge cemented out is separated from the solutionto leave a pure zinc chloride complex ion solution suitable forproducing high purity zinc oxide or zinc metal. Methods availableinclude direct precipitation of zinc oxychloride/zinc oxide/zinchydroxide. The alternative method of liquid-liquid extraction forchloride complex extraction may also be used.

After displacement by zinc of the other metals in the metal chloridecomplex, heavy metal cementation tank 24 feeds the zinc chloride complexZnCl₃ ⁻, preferably with a specific gravity of 1.43, in stream 700 toSimonkolleite precipitation tank 26. The zinc chloride complex in stream700 may have a purity of 99%+. Tank 26 also receives recycled water instream 560 in a first precipitation step and lime in stream 9100 duringa second precipitation step. As water in stream 560 is added toprecipitation tank 26, Simonkolleite is formed and precipitates out ofsolution. Approximately 35% of the zinc will precipitate out asSimonkolleite by the addition of water. Lime (calcium hydroxide) is thenadded to the solution in stream 9100 which results in approximately 60%of the zinc precipitating out as Simonkolleite. These two precipitationsteps can be performed separately or simultaneously. The remainingapproximate 5% of the zinc will remain as a zinc chloride complex andwill be recycled through streams 750, 820, 850, 900 and 200 back intothe process. Accordingly, there is an approximate 95% recovery rate ofthe zinc during the precipitation step in tank 26 during each pass.Specifically, at a temperature of 25 degrees Celsius, stream 750comprises a slurry of solids and solution. The solids portion comprisesapproximately 95% Simonkolleite, and approximately 5% zincoxide/hydroxide. The aqueous solution portion comprises approximately500 ppm of zinc in a zinc chloride complex form, as well asapproximately 42% calcium chloride leach solution and potassium andsodium chloride at their solubility limits (usually less than 5%).

The production of Simonkolleite, a zinc oxychloride, (and depending onconditions, small amounts of zinc oxide and zinc hydroxide) ispredominately produced by the following reactions, which occursimultaneously in tank 26:5Ca⁺⁺ 2+ZnCl ₃ ⁻+8Ca(OH)₂Zn₅Cl₂O₄.5H₂O+13CaCl₂  Equation 7 Ca⁺⁺2ZnCl₃ ⁻+2Ca(OH)₂2ZnO+3CaCl₂₊₂H₂O,  Equation 8ZnCl₂+Ca(OH)₂ZnO+CaCl₂+H₂O,  Equation 9wherein the Simonkolleite is Zn₅Cl₂O₄.5H₂O.

Prior to the zinc chloride complex being fed to tank 26, stream 700 iscooled to a temperature of approximately 25° C., down from a temperatureof approximately 90° C., because at these lower temperatures, moreSimonkolleite is formed from the zinc chloride complex. For example, at90° C. only approximately 40% of the zinc chloride complex isprecipitated as Simonkolleite. The reaction of the zinc chloride complexto Simonkolleite is exothermic so tank 26 must also be cooled tomaintain the tank at 25° C. during the reaction. As stated above,precipitating the zinc chloride complex at lower temperaturesunexpectedly yields high purity rates of the final product zinc oxidebecause the intermediary (Simonkolleite) is formed at high purity ratesat the cooler reaction temperatures, because the precipitation ofmagnesium (from the commercial calcium hydroxide) does notco-precipitate at such high rates compared to precipitation rates athigher temperatures.

Precipitation tank 26 feeds approximately 95% pure Simonkolleitecombined with small amounts of zinc oxide and zinc hydroxide solids,along with approximately 5% zinc chloride complex, in stream 750 to aSimonkolleite filter 28. Filter 28 also uses recycled water wash instream 540. There are two techniques for removing the sodium, orpotassium chloride as the case may be, which may crystallize out withthe Simonkolleite precipitate. A first preferred technique comprisespassing a hot filtrate, stream 800, through the oxychloride cake todirectly dissolve the chloride salts. The second, is to use hot water instream 540 in a counter current flow washing circuit to wash theSimonkolleite to cause the sodium or potassium chloride to solublize outof the Simonkolleite cake so that it does not flow with theSimonkolleite in stream 1000. In the second case, the countercurrentwash solution is chilled to 20° C. and the potassium chloride iscrystallized and filtered out, while the solution is then returned totank 26.

In particular, during the preferred technique, the Simonkolleite ispressed in filter 28 so that the liquid is pressed from theSimonkolleite precipitate and leaves the filter in stream 800. Theliquid in stream 800 comprises the unprecipitated 5% zinc chloridecomplex, and the calcium, potassium and sodium chlorides. A smallportion of stream 800 is passed to filtrate tank 34, is heated from 25°C. to approximately 80 to 90° C., and is then fed via stream 810 to washthe Simonkolleite filtrate cake. This wash removes any potassium orsodium chloride that may have precipitated earlier due to the coldertemperatures of the filtrate in filter 28. The Simonkolleite cake isthen pressed again in the filter. The liquid solution of this pressleaves the filter as stream 820. The remainder of stream 800 which wasnot fed to filtrate tank 34, and is designated as stream 812, iscombined with stream 820 and is fed to evaporator 32. Accordingly,stream 820 comprises the calcium chloride solution, a small amount ofunreacted zinc chloride complex, the potassium and/or sodium chloridesand dilution and wash water, and typically is at a temperature ofbetween 80 to 90° C.

The Simonkolleite cake, which is now free of potassium and sodiumchlorides, is then washed with water in stream 540 to remove solublechlorides, such as calcium chloride. Stream 540 typically comprisesfresh water or recycled water that already includes small amounts ofcalcium chloride, and is at a temperature of approximately 80° C. TheSimonkolleite is pressed so that stream 550 exiting filter 28 compriseswater at approximately 80° C. and some calcium chlorides. Stream 550 issent to a spent Simonkolleite wash tank 30 which feeds the dilutionwater in stream 560 to the Simonkolleite precipitation tank 26 and instream 570 to heavy metal cementation tank 24.

Still referring to filter 28 of FIG. 1, filter 28 feeds stream 820,which comprises the unprecipitated zinc chloride complex and thepotassium and/or sodium chlorides, to an evaporator 32. Stream 820 mayalso comprise small amounts of metal and calcium chlorides. Evaporator32 preferably comprises a crystallizer. Accordingly, salt crystals ofthe sodium chloride and potassium chloride crystallize out in theevaporator as water is evaporated. The concentration of calcium chlorideafter evaporation is sufficiently high to cause excess sodium andpotassium chloride entering the process to be crystallized when theirsaturation level has been met. The evaporated water is fed in stream 510for reuse in zinc oxide washing at filter 38 and in stream 500 forwashing at filter 16. A filter 40 or a centrifuge or other liquid-solidseparation method may be used to remove the crystals, which are thendried. The crystals may be blended in proper ratio with Cryolite to makea commercial grade aluminum salt flux, or the crystals may be fed viastream 1700 to a manufacturer of aluminum salt flux for further blendingand processing. A solution leaves filter 40 as stream 900 and typicallycomprises a high concentration leach solution of calcium chloridesuitable for recycle. The solution generally has a composition of 34%chloride typically comprising 53% calcium chloride or over 50% calciumchloride, and less than 3% potassium chloride, and less than 1% sodiumchloride. The solution also includes small amounts of zinc, magnesiumand other metals as chloride complexes which did not get removed inprior steps.

Stream 900 is fed to calcium chloride storage tank 22. A portion of thecalcium chloride may be sold to an outside source via stream 1800 tomaintain system chloride balance. A majority of the calcium chloride maybe recycled via stream 200 to leach reactor 20.

Referring to the salt separation and recycle portion of the process, thediluted and barren leach solution is recycled by the evaporation ofwater from the filtrate of the zinc oxychloride precipitation step. TheNaCl, KCl, and CaCl₂ can be crystallized out when the correctconcentration of CaCl₂ to water is reached for the next leach. Bydropping the temperature below 90° C. for a portion of the CaCl₂concentrate, the concentration can be controlled by crystallizing outexcess CaCl₂ which builds up in the system due to the entry of chloridesin the feed. These crystals can be dried for sale or sold as a solution.The barren leach solution is regenerated by the evaporating of wateruntil the desired calcium chloride concentration for the next leachsolution is reached. The temperature typically is held between 70° C.and 130° C. and may be removed by heat or vacuum or a combinationthereof. The water is removed such that the specific gravity is between1.49 and 1.55. The sodium, potassium and other salts which precipitateduring evaporation are filtered out. The salts are dried and aresuitable for use in the Aluminum Smelting Process as a flux. The NaCland KCl can also be recycled as feed stock. The NaCl and KCl may also beused as the feed stock to a chlor-alkali plant producing sodium andpotassium hydroxide and hydrochloric acid which can be used in theprocess and the excess sold commercially.

In one embodiment, caustic, also referred to as sodium hydroxide, isintroduced into tank 26 via stream 9100 (as will be described withrespect to FIG. 2) to allow a sufficient amount of sodium chlorideproduction that will produce more aluminum flux. This option willmaintain the system chloride balance and avoid the production of calciumchloride at stream 1800. Accordingly, by adding a sufficient amount ofcaustic to reactor 26 the system can be regulated so that no calciumchlorides are produced thereby eliminating the need to dispose of excesscalcium chlorides in stream 1800 which does not have a market.

Referring again to the Simonkolleite portion of the process, the processcomprises the production of high purity Simonkolleite in equilibriumwith zinc oxide and other minor zinc oxychlorides from a concentratedchloride pregnant solution containing zinc chloride complex ions whichis of sufficient purity for subsequent conversion as a feed stock toproduce high purity zinc or zinc oxide. An unexpected method for theproduction of Simonkolleite (an intermediate oxychloride to be used inthe production of zinc oxide) or zinc metal, from a high concentrationchloride solution, first includes its partial precipitation through theaddition of water to a pregnant solution to reduce the chlorideconcentration and second, further precipitation through the addition ofa base. Using this process, 5% to 100% purity Simonkolleite can beproduced in equilibrium with zinc oxide. The solution may comprise oneor a combination of the following: calcium chloride, sodium chloride,potassium chloride or any other Group I and Group 2 chlorides. Thesolution typically contains zinc at a concentration greater than 500 ppmand preferably 10,000 ppm or greater. In a first stage of the process,dilution water is added to the solution such that up to 35% of the zincunexpectantly precipitates as Simonkolleite. This unexpectedprecipitation conserves the amount of base to be added and thus reducesthe base required by 35%. The specific gravity is reduced to below 1.43(1.39 preferred) or the chloride concentration is reduced by othermethods to an equivalent level. A base is added in a second step to thesolution to precipitate the majority of the remaining zinc chloridecomplex present in the original solution as Simonkolleite. The base maycomprise calcium hydroxide or calcium oxide. The base may also compriseone or a combination of sodium hydroxide or potassium hydroxide. Thefirst and second steps may be combined into a single step. The processis conducted at atmospheric pressure and the temperature preferably isbetween the solution crystallization point and the solution boilingpoint. The addition of water during the first stage is exothermic. Theaddition of a base to the solution is also exothermic. The solution ismixed to facilitate the reaction. The solids are separated from thesolution after precipitation. The solids are then washed and containless than an expected amount of lead and cadmium which remained in traceor very small quantities after the metal removal. The solution pH istypically below 9.0. The solution pH may be held at or below 7.2 tominimize the co-precipitation of magnesium present in commerciallyavailable calcium oxide and hydroxides. Washing with a hot solution ofcalcium chloride in stream 810 removes excess potassium chloride. Thesolution of calcium chloride in water typically is held at a temperaturebetween 70-130° C. and has a specific gravity of between 1.41 to 1.45 at90° C. The wash solution may be recycled by separating the solids fromsolution. Washing the solids with water removes excess soluble chloride.The wash water temperature typically is between 20-100° C. The filtratetypically is suitable for regeneration and reuse. The resultant productis an ideal feed stock for producing small particle and high surfacearea zinc oxide, known as active or reactive zinc oxide.

Referring now to the Metals Loop of the process, heavy metal cement tank24 feeds the displaced metals in stream 2000 to a cadmium leach tank 42.Leach tank 42 also receives hydrochloric acid in stream 9200 from ahydrochloric acid storage tank 44. Leach tank 42 feeds the resultingsolution in stream 2100 to a lead filter 46. Lead filter 46 feeds theseparated lead in stream 2150 to a lead furnace 50 and the displacedmetals solution, without lead, in stream 2200 to a cadmium cementationtank 48. Lead furnace 50 feeds lead in stream 1400 for transport tooutside sales. Cadmium cementation tank 48 receives zinc powder instream 9300 from a zinc powder storage bin 52 and feeds the resultingsolution in stream 2300 to a cadmium filter 54. Cadmium filter 54 feedsthe separated cadmium solution in stream 2350 to a cadmium furnace 56and water, including some zinc chlorides, in stream 2500 to heavy metalcement tank 24. Cadmium furnace 56 feeds the cadmium in stream 1500 fortransportation to outside sales. This completes the Metals Loop, whichshould be understood by those skilled in the art.

Referring now to the Zinc Oxide Loop of the process, the cleanSimonkolleite cake, with a very low chloride concentration, is fed instream 1000 to reslurry tank 36, along with water in stream 530.Re-slurry tank 36 feeds the watery Simonkolleite slurry in stream 1100to a zinc oxide reactor 60. The contents of reactor 60 are heated to atemperature of approximately 160 to 190° C., and preferably to atemperature of approximately 170° C. (The reactor generally is heated toa temperature of approximately 140-150° C. when caustic is added toreactor 60, as discussed with respect to FIG. 2). Reactor 60 alsoreceives lime (calcium oxide) in stream 9150 from a lime storage stank62. The temperature of the solution, and the addition of lime to thesolution, results in the conversion of the Simonkolleite to high puritygrade zinc oxide by the following reaction:ZnCl₂.4Zn(OH)₂.H₂O+Ca(OH)₂5ZnO+6H₂O+CaCl₂,  Equation 10wherein the Simonkolleite can be written as ZnCl₂.4Zn(OH)₂.H₂O or asZn₅Cl₂O₄.5H₂O.

The resulting zinc oxide, calcium chlorides and water are fed in stream1150 to a zinc oxide product filter 38, which also receives recycledwater in stream 510 from evaporator 32. Filter 38 typically is atwo-stage filtration process. First, the filter presses the solution sothat the solid zinc oxide precipitate remains in the filter and thecalcium chlorides and water leave the filter in stream 580 to filtratetank 68. Water in stream 510 is then used to wash the zinc oxide solidsso that the small amount remaining of the calcium chloride and the wateris removed in stream 520 to spent zinc oxide wash tank 64. The zincoxide solids are then removed from the filter in stream 1200 to a zincoxide dryer and pulverizer 66. Stream 530, comprising a very dilutesolution of calcium chloride and water, is recycled back to reslurrytank 530 and stream 580, comprising dilute calcium chloride and water,is recycled back to filter 28 via stream 540.

Dryer 66 typically is operated at 125 to 500° C., and at a preferredtemperature of 200° C., to remove moisture and hydrates, which compriseapproximately 3% of the stream, from the zinc oxide. The dryer feeds thedried product in stream 1300 to outside sales. Stream 1300 typicallycomprises 98% zinc oxide, but, under preferred reaction conditions,comprises approximately 99% zinc oxide. As stated above, the zinc oxideproduced is active and has a large surface area, rendering the zincoxide suitable for use in rubber and plastic compounding. This completesthe Zinc Oxide Loop. This process could also be used for otheroxychlorides such as the recovery of copper oxides. For example, oncecopper oxychloride is made, this process can be used to make copperoxide.

The process of the present invention converts Simonkolleite and otherzinc oxychlorides to zinc oxide, meeting specifications and suitabilityfor use as rubber grade zinc oxide with more than 99% pure zinc oxide.The zinc oxide has less than 2,000 ppm chlorides and under preferredconditions less than 1000 ppm and has a particle size of 0.05 micron to0.5 micron. The process produces an active zinc oxide with a surfacearea of 10-70 m/gram, providing better reactivity and economy comparedto French processed zinc oxide. This large amount of surface area offersa more reactive product allowing its use at lower levels than iscurrently practiced.

The precipitate, Simonkolleite in equilibrium with zinc oxide, zinchydroxide, and/or other zinc oxychlorides and their hydrates, producedfrom zinc chloride complex solutions high in chlorides or from theaddition of a base to concentrated zinc chloride solution of sufficientpurity, or from other sources, is used as the feed stock to make highpurity zinc oxide. The solids of the equilibrium solution are separatedfrom the high chloride solution from which they were produced byfiltration or other methods. The solids may be washed to further reduceresidual soluble chlorides (not oxychlorides). The solids are thenreslurried/repulped in fresh water or recycled water containing lessthan 5% chlorides. Calcium hydroxide or oxide is added to the reslurriedsolution of the Simonkolleite/zinc oxychloride/zinc oxide instoichiometric quantities or greater to be available to react with thechloride ions in the Simonkolleite and zinc oxychloride. Maintaining thepH between 6.9 and 7.4 keeps the Mg in the lime from precipitating withthe zinc oxide. Alternatively, sodium or potassium hydroxide is added tothe solution in stoichiometric quantities or greater to react with thechloride ions in the Simonkolleite and zinc oxychloride. The sodium orpotassium hydroxide can be added in significant excess to reduce thechloride content below 500 ppm. The re-pulped slurry is processed in anautoclave or a reactor with a mixing capability. The temperature for thereaction typically is 140 to 200° C., and preferably 150° C. when usingcaustic or 170° C., when using calcium hydroxide for one hour at thenatural pressure of the boiling solution. The crystal structure of thezinc oxide is tiny needle like crystals compared to prism and/or pancakeshapes. The zinc oxide surface area typically is in excess of 20 m/g andless than 70 m/g. The particle size typically is between 0.1 and 0.5microns. The lead, copper and magnesium contaminants are low. Inparticular, the lead level of the zinc oxide is below 60 ppm. Thecadmium level of the zinc oxide is below 30 ppm. The magnesium contentof the zinc oxide level is below 300 ppm. The zinc oxide cake is washeduntil free of soluble chloride salts produced from the reaction. Thefinal chloride content typically is below 1000 ppm. The filter cake isdried to a temperature between 125° C.-600° C. The product can beproduced as a dust free pellet or a briquette without binder. Thesequence of process steps set forth above can be reconfigured tooptimize recovery selection and to reflect changing economics.Additional cementation steps utilizing more than one dust forcementation may be added in sequence as required for additional orselective metal removals.

As described above, the process is a batch process. However, the processof the present invention may also be carried out as a steady statecontinuous process.

In particular, the process works based on the relative insolubility ofiron, magnesium, calcium and others under conditions of pH 6 and abovein a high calcium chloride concentration, while the desirable metalswill exchange with the calcium, becoming metal chlorides andprecipitating calcium hydroxide driving the reaction in reverse. Theoverall efficiency of the process depends on the chloride and calciumconcentration, the temperature, pulp density, reaction residence time,and mixing rate.

The process provides conditions to remove halogen salts. Fluorides inthe feed are converted to insoluble calcium fluoride removing them fromfurther reaction.

The process is environmentally sound with the opportunity of no liquidwaste discharge and no air/odor emissions. The leach process, whileoperating at atmospheric pressures, does not emit off gases and thusposes no odor emission issues. The process is especially attractive as areplacement for sulfur leach chemistry and roasting used extensively inthe mining industry. It is especially attractive as a means to avoidwaste disposal landfilling. A typical feed stock would be Electric ArcFurnace Flue dust and its subsequent pyrolysis derivatives dusts whichgenerally contain chloride salts formed in the thermal decomposition ofPVC from scrap cars, in the presence metal vapor and metal oxides.

Still referring to FIG. 1, now that a preferred embodiment of the basichydrometallurgical portion of the process has been described, a detaileddescription of the hydrometallurgical pilot plant process will bedescribed. The scale of the pilot plant process is merely increased forcommercial production.

I. Material Preparation

-   -   A. Feed material: may be chosen from the following list:        -   1. Zinc Hearth Dust: the zinc/metals concentrated dust            derived from the off gasses of processing KO61 Electric Arc            Furnace Dust in combination with coal or coke in a reduction            furnace to remove volatile metals and salts and reduce iron            oxide and other metals to their metallized form (stream            9950).        -   2. KO61 EAF Dust: the dust derived from the off gasses of            the Electric Arc Furnace processing of scrap steel            containing approximately 20-24% zinc (stream 9900)        -   3. Other metal feed stocks including one or more of the            following: zinc, lead, cadmium, silver, copper, iron,            magnesium, halogen salts and other heavy metals; metal            furnace dusts, smelting dusts, waste sludges, mill tailings,            ores which containing metal oxides, metal hydroxides, metal            ferrites, metal sulfates/sulfites/sulfur compounds,            carbonates, metal bearing materials containing chlorides or            fluorides; zinc oxide dust (zinc concentrate) recovered from            pyrolysis/furnace operations to roast or reduce (metallize)            metal ore or waste metal bearing materials such as Electric            Arc Furnace Dust when making direct reduced iron; Electric            Arc Furnace Flue dust KO61 derived from the off gasses of            the Electric Arc Furnace processing of scrap steel where            said dust contains approximately 10-40% zinc; and Electric            Arc Furnace dust containing zinc ferrite.    -   B. Leach Solution Preparation        -   1. Prepare the initial or make-up “Leach Salt Solution”            solution.            -   a) Add 69.76 kg (153.79 lb.) water into the CaCl₂ tank                (tank 22). Record the weight and the time of addition.            -   b) Add 77.44 kg (170.73 lb.) CaCl₂ to the water in tank                22. Record the weight, time and temperature of addition.        -   2. Mix the solution vigorously for 10 minutes or until all            CaCl₂ is dissolved.        -   3. Determine the specific gravity using a hydrometer at            90° C. The target specific gravity is 1.49.    -   C. Ca(OH)₂ Solution Preparation        -   1. Prepare a solution containing 18.9% CaO (should be            prepared 24 hours before intended use).            -   a) Place 50 kg of water inside the CaO tank (tank 62).            -   b) Add 11.65 kg of CaO to the CaO tank.            -   c) Allow the solution to mix for at least 30 minutes            -   d) Wait at least 2 hours before using the solution    -   D. Process Configuration        -   1. FIG. 1 shows a schematic flow diagram of the zinc oxide            (and other products) recovery process utilizing a pyrolysis            furnace to make an iron/calcium briquette, and crude hearth            dust leached with calcium chloride; and utilizing lime for            zinc precipitation (the preferred embodiment described            below);        -   2. FIG. 2 shows a schematic flow diagram of the zinc oxide            (and other products) recovery process showing a pyrolysis            furnace to make an iron/calcium briquette, and crude hearth            dust; and utilizing caustic for zinc precipitation and then            its recycle utilizing a chlor-alkali-plant.        -   3. FIG. 3 shows a schematic flow diagram of the zinc oxide            recovery process directly from K061 utilizing lime for zinc            precipitation; and using a reduction furnace produce to            metallized iron briquettes        -   4. FIG. 4 shows a schematic flow diagram of the zinc oxide            recovery process from K061 utilizing sodium hydroxide for            zinc precipitation; and using a reduction furnace to produce            metallized iron briquettes.

II. Standard Pilot Plant Operating Procedure-Pyrometallurgical

Prior Art: Reduction processes for metallizing (converting iron oxide toiron metal) iron bearing wastes and specifically KO-61 have used onlycarbon (coke, coal, etc.) to reduce the metals and allow thevaporization of non-ferrous metals like zinc. Commercial recovery ofKO-61 is general inefficient, removing only 80-90% of the zinc andmetallizing only 90% of the iron.

Reduction/Metallization/Non-Ferrous metals removal:

Calcium hydroxide, calcium carbonate, calcium oxide, magnesiumhydroxide, magnesium carbonate, or magnesium oxide along withmetallurgical coke is mixed with KO61 (or other iron and non-ferrousbearing streams) and a small amount of water. The blend is often madeinto a “green ball” being charged to a metallizing reductionfurnace/rotary hearth furnace, rotary or stationary kiln or othersimilar furnace process which significantly increases the recovery(removal) of such non-ferrous metals, including, but not limited to,zinc, lead, silver and cadmium. The furnace is further fired withnatural gas or other gas, liquid or solid fuel to achieve the necessarytemperatures. The addition of the calcium/magnesium improves the amountof iron metallization occurring in the product and besides providing animproved source of iron it can provide a balanced source of calcium andmagnesium flux which is required in the iron and steel making process toprovide for slag and steel chemistry requirements. Normally this type ofmetallization process produces DRI or sinter which has about 90% ofthese wanted non-ferrous metals removed. A typical product produced fromKO-61 contains approximately 5% (50,000 mg/kg) zinc, while a briquettemade from the inventive process contains between 600-900 mg/kg (seepilot plant results in Table 17). The addition of the calcium hydroxideto the KO-61 unexpectedly improves the removal to over 99% for zinc andlead removal, and over 95% for cadmium (see pilot plant results Table18). This results in a preferred product being sold to the steel millsince the final fluxed sinter briquette product contains very low levelsof non-ferrous metals when compared to other iron materials producedfrom KO-61 that do not use the addition of calcium and magnesiumhydroxide/oxide. The process has the added advantage of providing aneeded source of calcium and magnesium oxide which is required in theprocessing of iron ore and scrap to iron and steel. The fluxed sinter isbriquetted or pelletized under elevated pressure in a standardbriquetting machine so that it is not friable and will withstand therigors of being delivered into the Electric Arc Furnace or other steelfurnace with less than 1% being less than ¼ inch in size. This offersthe mill the added advantage of being able to convey needed calcium andmagnesium into the furnace as part of the iron briquette, using standardscrap handling magnets, just as is done for steel scrap, with the samehandling costs. Because of the large size of the briquettes and theirstructural strength they can also be delivered to the steel melt pot, aslarge chunks, pneumatically or by crane magnet without being lost in theflue gas streams out of the steel furnace.

Tests were conducted in a laboratory scale using a small tube furnace tosimulate the conditions in a full scale reduction furnace. In testingRound 1, approximately 33 gm of calcium hydroxide was added to 100 gmKO-61 with low levels of zinc and other non-ferrous metals and wascompared to KO-61 processed without calcium hydroxide. The calcium addedproduct consistently had 0.1-0.5% zinc, while under the same conditions,zinc ranged from 4.6-5.8% in the final product when not adding calciumoxide/hydroxide. In Round 2 tests, calcium hydroxide was added at 18 gm,to coke at 15.7 gm and 100 gm KO-61 with a small amount (approximately10%) of water and formed into a ball. The “green” ball was preheated for5 minutes at 275° C. and processed at 1030-1100° C. for 15-65 minutes.Zinc levels under the best conditions were in the range of 0.06-0.09%.Best results were obtained at operating temperatures of 1030° C. to1070° C. and retention times of 30-65 minutes. For one knowledgeable inthe art of metallization a much wider range of calcium hydroxideaddition is possible as is the variance of amounts of carbon to achievethe desired stoichiometry in the furnace, based on the exact feedcomposition. The preferred conditions are a furnace retention time of 45minutes at 1050° C. with 18 to 33 (but a wider range is likely) parts ofdried calcium hydroxide, 15.7 parts of low sulfur metallurgical coke,and 100 parts Electric Arc Furnace Flue dust KO-61. It is believed thatthe calcium serves to increase the reaction activity of the chemicalprocesses occurring in the process.

III. Standard Pilot Plant Operating Procedure-Hydrometallurgical

Prior Art: Leaches performed with chloride solutions have historicallybeen performed with a chloride salt and then the addition of acid or ametal acid salt such as ferric or cupric chloride. These approachesrequire the addition of a costly additive or they impart unwanted ionsto the solution and carry unwanted ions in the pregnant solutions to thenext processing step requiring further purification or producinginferior product quality. Those leaches done with calcium chloridesolutions like U.S. Pat. No. 5,078,786 (specifically limited to jarositecontaining materials) requires a low pH, and leach pressures requiring apressure vessel. Other processes use sodium chloride with thedisadvantage of requiring acid or chlorine, resulting in increasedchemical costs and the generation of iron chlorides needing furthertreatment and disposal.

Leach:

Chemistry: Probable leach reactions for Zn, Cd, Cu and Pb (hereinrepresented by the reference M) are set forth below as equations 3-5.Other elements like Fe, Mg, Ca, Cr and Mn do not generally leach underthe stated concentration unless the chloride concentrations aresubstantially reduced. The reactions are as follows:3CaCl₂+2MO+2H₂O Ca⁺⁺+2MCl₃ ⁻+2Ca(OH)₂  Equation 3CaCl₂+MCl₂[Ca⁺⁺+MCl₄ ⁻]  Equation 4CaCl₂+MO+H₂O MCl₂+Ca(OH)₂,  Equation 5wherein M=Zn, Cd, Cu and/or Pb, and wherein the metal chloride complexis MCl₃ ⁻.

The process is unique and different than alternative processes in thatthe equilibrium of extremely high calcium chloride concentrations drivesthe reactions “backwards” forcing the metals into solution as chloridecomplexes, while the calcium precipitates as calcium hydroxide. By“backwards” applicants mean that the reaction is driven to unexpectedlyproduce calcium hydroxide and the chloride complex. When fluorides arepresent in the feed insoluble calcium fluoride is formed. One of thebeneficial aspects of the process is that iron is essentially insolubleat these concentrations and is thus not carried to subsequentoperations, or if it is, the quantities are sufficiently small to beeasily removed and recycled without making large quantities of ferrichydroxide for disposal.

While the leach temperature may be above 80° C., or the crystallizationpoint of the solution, a temperature just below the boiling point ispreferred. The preferred reaction pressure is ambient to control capitalequipment costs, but the process can take place at more elevatedtemperatures and pressures.

A. Metals Leach Reaction

-   -   1. Add zinc hearth dust to the reactor 20.        -   a) Obtain a sample of CaCl₂—H₂O solution to be added to            reactor 20. (Label: Leach Salt Solution from CaCl₂ tank 22).        -   b) Measure out 61 L of the CaCl₂—H₂O solution and transfer            it to reactor 20 (stream 200).        -   c) Measure out 2.54 kg zinc hearth dust from storage tank            18. (stream 100). Record the weight.        -   d) Add this material to the 61 L of CaCl₂—H₂O solution at            90° C. at atmospheric pressure. Record the time and            temperature of addition.    -   2. Latch and tighten the bolts around the cover of reactor 20.        This will prevent any water evaporation from the reactor during        the leaching process.    -   3. Heat the leach reactor 20 to a reaction temperature of        120° C. The reaction time at this temperature is approximately        ninety minutes at 0 psig.    -   4. Every 30 minutes, monitor the temperature, and pressure, and        note any fluctuations.    -   5. After 90 minutes at the reaction temperature, remove a 125 ml        sample of the slurry.        -   a) Record the time and the reactor temperature        -   b) Measure the pH, mV, and temperature of the sample in the            Lab, conduct Inductively Coupled Plasma Emission            Spectrometer (ICP) analysis.

B. Pre-Heat Filter

By heating the filter before use, this step insures that the calcium andmetal salts do not crystallize out during filtration.

-   -   1. Place a clean filter leach cake cloth into filter 16. Make        sure the cloth is centered before closing the filter.    -   2. Push down on the end labeled “pump” on the hydraulic pump        located on the filter table. Run the hydraulic pump until the        pressure reading on the pressure gauge is close to the red line        which indicates a pressure of 2000 psi. DO NOT pump past the red        line.    -   3. Close the valve located behind the pressure gauge of the        hydraulic pump. This will help to prevent leaks through the        filter.    -   4. Place insulation over the filter and tighten the wires around        it    -   5. Place the heat gun at the bottom of the filter, pointing up        towards the filter. The heat gun, as well as all of the other        pre-heating steps, will help to prevent CaCl₂ precipitation in        the leach cake and to speed up the rate of filtration.    -   6. Turn on the hot water pump switch (stream 500).    -   7. Slowly turn the hot water valve open and allow the water to        flow through filter 16 for one minute.    -   8. Close the water valve and open the air valve to blow out        excess water in the filter.    -   9. Close the air valve.    -   10. Attach the hose exiting reactor 20 to the hose entering the        filter pump.    -   11. Open the air bleed valve located to the left of the reactor        pressure gauge to release any vapor pressure that had built up        inside the reactor.    -   12. Unlatch the reactor door latches and insert the outlet        recirculation hose (not shown) of filter 16 into the reactor.        Secure the hose with plastic cable latches.    -   13. Make sure all the valves on the filter are closed.    -   14. Open the reactor outlet valve located at the bottom of the        reactor.    -   15. Turn on the pump to the filter.    -   16. Slowly open the valve to the outlet recirculation hose. This        allows the hot slurry to circulate from the filter pump line and        back into the reactor. The slurry will heat up the pump line as        it circulates.    -   17. Allow the slurry to circulate in the lines for a couple of        minutes before proceeding to the Leach Slurry Filtration step.    -   18. Check the temperature of the solution in CaCl₂ leach        solution tank (tank 22) to make sure is approximately 90° C. and        the specific gravity of the solution is at 1.49. Add water if        necessary to adjust the specific gravity to 1.49.    -   19. Take a CaCl₂ Leach Salt Wash solution sample. (Label: Leach        Salt Wash (before wash))        -   a) Record the time and temperature of the solution.        -   b) Measure the pH, mV, and temperature of the sample in the            lab.        -   c) Conduct an ICP analysis.

C. Leach Slurry Filtration

-   -   1. Maintain the solution temperature between 90°-1101° C. while        filtering through the Larox filter 16.    -   2. Close the outlet recirculation valve slightly to decrease the        amount of solution that flows back into reactor 20.    -   3. Slowly open the valve to the filter and allow the slurry        pressure to reach a pressure within the range of 60-90 psi.    -   4. Collect the filtrate in buckets (stream 625).    -   5. Close the filter outlet recirculation valve.    -   6. Check the level of solution inside the reactor and allow the        pump to continue pumping until the sound of the pump starts to        change.    -   7. Close the reactor outlet discharge valve.    -   8. Open the valve to the Larox filter diaphragm pump.    -   9. Open the valve located on the left side of the pressure gauge        on the Larox filter diaphragm pump. This allows water to be        pumped into the diaphragm and press the solids inside the filter        to “squeeze” more fluids out of the cake.    -   10. Slowly open the filter outlet valve and pump the remaining        slurry in the lines into a bucket (stream 650).    -   11. Turn off the pump.    -   12. Detach the hose and attach it to the outlet located at the        bottom of CaCl₂ tank 22.    -   13. Transfer filtrate to the cementation Tank 24 after        filtration and hold at approximately 90° C.+.    -   14. Determine the filtrate volume.    -   15. Determine the specific gravity of the filtrate.    -   16. Remove a 50 ml sample of the filtrate. (Label: Leach        Filtrate)        -   a) Record the time and reactor temperature.        -   b) Measure the pH, mV, and temperature of the sample in the            lab.        -   c) Conduct an ICP analysis.        -   d) Note any observations.    -   17. Measure and record the volume and the weight of the        remaining slurry.    -   18. Dispose of the remaining slurry properly. (Note: the        remaining slurry is disposed of during this pilot plant process        because only small quantities are used for measurement purposes.        During commercial processing all of the products and wastes of        the process are fed back into the system so that all streams are        recycled.)

D. “Leach Salt Solution” Wash

This step is designed to remove the entrained pregnant leach solutionfrom the filter cake and has been found to further increase metalsrecovery.

-   -   1. Open the valve located at the bottom of the CaCl₂ tank 22.    -   2. Turn on the pump to filter 16.    -   3. Close the air valve (hereinafter the yellow lever) to        pressurize the Larox filter diaphragm pump.    -   4. Slowly open the valve to the filter and close the valve to        the diaphragm return/relief line (hereinafter the blue lever) on        the Larox filter diaphragm pump to relieve the pressure in the        diaphragm.    -   5. Wash the filter cake solids from filter 16 with the CaCl₂        solution from tank 22 (total of 13.6 L) solution. (stream 210)        -   a) Close the filter outlet valve to stop the flow of            filtrate when 3 L of CaCl₂ solution has been collected. This            will allow the solution to fill the filter and soak the            solids to achieve maximum zinc recovery.        -   b) Allow the solids to soak for two minutes.        -   c) Open the filter outlet valve.        -   d) Wash the solids with an additional 10.6 L of CaCl₂            solution.    -   6. Remove 50 ml of spent CaCl₂ wash solution (after washing the        solids) (Label: Spent Leach Wash Salt)        -   a) Record the time and temperature of the wash solution        -   b) Measure the pH, mV, and temperature of the sample in the            lab.        -   c) Conduct an ICP analysis.    -   7. Open the valve (yellow lever) to the Larox filter diaphragm        pump.    -   8. Open the valve (blue lever) located on the left side of the        pressure gauge on the Larox filter diaphragm pump. This allows        water to be pumped into the diaphragm and press the solids        inside the filter to “squeeze” more fluids out of the cake.    -   9. Detach the feed hose, stream 200, (hereinafter the green        hose) from the filter.    -   10. Slowly open the outlet valve and pump the remaining solution        in the lines into a bucket and transfer it back into the CaCl₂        tank 22.    -   11. Turn off the pump.    -   12. Determine the total spent CaCl₂ wash volume and specific        gravity.    -   13. Add spent wash solution to the leach filtrate in cementation        Tank 24 and maintain the temperature in the range of 90-110° C.        (stream 650).

E. Water Wash

This step is included to wash out and recover chlorides, recoverpotassium chloride and sodium chlorides, and increase non-ferrous metaland specific copper recovery. It has been found that as the chlorideconcentration slowly decreases that the solubility of copper increases.Wash time and duration of the contact directly effect the recovery andcan be varied to increase or decrease specific metals recovered based onthe various feed materials charged to the leach.

-   -   1 For this test, fresh water will be used, but recycled water        may also be used.    -   2. Turn on the water pump to the water tower.    -   3. Close the valve (yellow lever) on the Larox filter diaphragm        pump.    -   4. Slow open the water valve to the filter and close the valve        (blue lever) on the Larox filter diaphragm pump.    -   5. Use 6.3 L of H₂O to wash the leached cake solids (stream 500)        following the CaCl₂ wash (stream 210).        -   a) Close the filter outlet valve to stop the flow of            filtrate when 3 L of wash water has been collected. This            will allow the solution to fill the filter and soak the            solids to achieve maximum salt recovery.        -   b) Allow the solids to soak for two minutes.        -   c) Open the filter outlet valve.        -   d) Wash the solids with an additional 3.3 L of wash water.    -   6. Open the valve (yellow lever) to the Larox filter diaphragm        pump.    -   7. Open the valve (blue lever) located on the left side of the        pressure gauge on the Larox filter diaphragm pump. This allows        water to be pumped into the diaphragm and press the solids        inside filter 16 to “squeeze” more fluids out of the cake.    -   8. Determine the spent wash water volume. Theoretical volume=6.3        L    -   9. Obtain a 50 ml wash water sample (after washing the solids,        Label: Spent Leach Wash Water (after wash)).        -   a) Record the time and temperature of the wash water            solution sample.        -   b) Measure the pH.        -   c) Conduct an ICP analysis.    -   10. Add the spent wash water to cementation Tank 24 (stream        600).    -   11. Obtain a sample of a combination solution of leach filtrate,        leach salt wash, and water wash (solution in the cementation        Tank) (stream 675, Label: Leach Blend (filtrate+Salt+H₂O Wash)).        -   a) Record the time and temperature of the combined solution.        -   b) Measure the pH.        -   c) Conduct an ICP analysis.    -   12. Close the valve (yellow lever) on the Larox filter diaphragm        pump.    -   13. Slow open the air valve to the filter and close the valve        (blue lever) on the Larox filter diaphragm pump.    -   14. Allow the air to flow through filter for 1 minute.    -   15. Close the air valve to filter.    -   16. Turn off the heat gun and remove the insulation.    -   17. Open the filter by pushing the end labeled “release” on the        hydraulic pump.    -   18. Carefully remove the filter cloth and the cake from the        filter.    -   19. Obtain a sample of cake (Label: Leach Cake) (stream 1600).        -   a) Record the total weight of the cake.        -   b) Record the thickness of the cake.        -   c) Conduct an ICP analysis.        -   d) Store the solids inside a plastic bag. Label the bag            properly.

Cementation Step:

Lead, copper, cadmium, silver, other valuable metals and trace amountsof iron are cemented out in this process. The uniqueness of the processrelies on varying the chloride concentration with water dilutionconcentration, instead of pH control with acid addition, to set up theconditions where the addition of zinc, aluminum, copper, or iron powderor dusts will selectively cement out the desired metal to be recovered.The uniqueness of the process is that at natural pH conditions of 6 pHor higher the desired metals cement out and, unexpectedly, metal oxidesor hydroxides are not produced. Thus, the metals made are suitable formetal smelters/refiner/processors, production of elemental metal orother processes seeking reduced (metallized) metals or direct sale.

The process produces a filtrate (pregnant zinc solution), which containszinc chloride complexes, which is pure and ideally suited for theproduction of zinc oxychloride and subsequently rubber grade 99% activezinc oxide. The preferred solution contains less than 2 ppm of lead,copper, cadmium, and iron.

The preferred method for processing of the pregnant leach solution fromzinc hearth dust or KO61 is to cement all the metals simultaneously withzinc dust and directly sell or process the recovered metals in anadditional step to extract cadmium from smelter feed. The process can beperformed with a single stage or continuous countercurrent reaction, buta two stage countercurrent process is preferred to increase theefficiency of the zinc dust addition to the preferred utilization of 1.2times stoichiometric.

The prior art processes require the adjustment of pH below pH 6 throughthe addition of costly acid and then the addition of a base tofacilitate the desired cementation without the formation of unwantedmetal hydroxides or oxides. In contrast, the present invention uses a pHof approximately six or greater and does not require an acid or a baseto facilitate the desired cementation step conditions.

F. Iron Removal Pretreatment/Post Treatment: This step may be a)omitted, b) performed before the cementation of metals, or c) performedafter cementation.

-   -   1. Mix the combined solution (pregnant leach filtrate, CaCl₂        spent wash and the water spent wash).    -   2. Add additional water to the solution as required to adjust        the preferred specific gravity to 1.43.    -   3. Sparge the solution gently with air or oxygen at 70-90° C.        until saturated and floc forms. Accordingly, the iron is        oxidized by the injection of oxygen or air and the iron is then        separated.    -   4. Filter out the iron bearing solids using poly-electrolyte        coagulant as required (such as Nalco 9812).    -   5. Analyze to confirm that the iron content of the filtrate is        below 6 ppm.

G. Cementation of Metals from the Pregnant Leach Solution: A two stagecounter current reaction is preferred.

Stage I:

Mix the combined solution (pregnant leach filtrate, CaCl₂ spent wash,and H₂O spent wash).

-   -   1. Add water to the combined leach solution as required to        adjust the specific gravity to 1.43.    -   2. Remove a 50 ml sample of solution (Label: Cementation        Solution Start).    -   3. Combine the mixed solutions with the partially reacted zinc        and metal solids settled from Stage II and mix vigorously for 45        minutes at 90° C.    -   4. The following reactions take place in cementation:        -   a) Zn+[Ca+2MCl₃]M[CaZnCl₃]        -   b) Zn+MCl₂M+ZnCl₂    -   5. Obtain a sample of solution (Label: Cement, Stage 1@ Time        ______)        -   a) Record the time and date.        -   b) Measure the pH.        -   c) Conduct an ICP analysis.    -   6. Check with lab personnel to determine the lead level in        solution.        -   a) If more than ˜25% of the lead that came into cementation            Stage I is left in solution, add 20 g of zinc dust to the            solution and mix for another 30 minutes (level of Pb should            be <400 mg/kg).            -   (1) Obtain a sample of solution (Label: Cement, Stage 1@                Time ______)            -   (2) Record the time and date.            -   (3) Measure the pH.            -   (4) Conduct an ICP analysis.        -   b) Check with lab personnel to determine the level of Pb            remaining in solution (the level of Pb should be <400            mg/kg).        -   c) Pump all of the solution and solids in cementation Tank            24 into buckets using the gray pump.        -   d) Allow the solids to settle inside the buckets.        -   e) Decant the solution back into cementation Tank 24 for            Stage II.        -   f) Collect the remaining solids into one bucket and decant.            The cementation solids are now ready for the heavy metals            refining step. (stream 2000).            -   (1) Obtain a sample of the solids (Label: cementation                Solids).            -   (2) Record the total weight of the solids.            -   (3) Conduct an ICP analysis.            -   (4) Store the solids inside a plastic bag. Label the bag                properly.

Stage II

-   -   7. Determine specific gravity of the solution from Stage I. Add        water as required to set the specific gravity to 1.43 (stream        570).    -   8. At 90° C., add 80 g Zn dust into cementation Tank 24. Record        the weight, time, and temperature of addition.    -   9. Mix vigorously for 30 minutes.    -   10. Obtain a sample of solution (Label: cementation Stage 2 @        Time ______).        -   a) Record the time and date.        -   b) Measure the pH.        -   c) Conduct an ICP analysis.    -   11. Check with Lab Personnel to make sure the level in solution        of Pb<5 mg/kg, Cd<2 mg/kg, and Cu<1 mg/kg. If any of the heavy        metals in solution is higher than the values specified above,        add an additional 20 g Zn dust and mix for another 30 minutes        before filtering. Repeat as necessary.    -   12. Filter the solution using a cartridge filter and leave all        the solids inside the tank. The solids will be used in Stage I        of the next incoming batch (stream 700).    -   13. Obtain a 50 ml sample of the filtrate (Label: cementation        Stage 2 Filtrate)        -   a) Record the time and date.        -   b) Observe the pH.        -   c) Conduct an ICP analysis.

Determine the Solution Volume and Specific Gravity.

Transfer the filtrate into Precipitation Tank 26.

Repeat steps A-F for another run before proceeding to step G to amass tofull runs for the next step.

Heavy Metals:

Because some lead smelters may prefer feed without cadmium, theinventive process takes the heavy metals (stream 2000) and reacts themwith hydrochloric acid (stream 9200) to selectively leach cadmium. Thelead, copper, and silver are filtered or separated out (stream 2150).The cadmium solution (stream 2200) is cemented with zinc dust (stream9300) after pH adjustment under conditions well known in the industryand separated from the barren solution (stream 2500) by filtration. Thebarren solution can then be recycled (stream 2500) back into theprocess. Other extractive and separation processes may also be used.

The heavy metal concentrate (stream 2150) is pelletized and may then beintroduced to a molten metal bath and made into ingots (stream 1400).The cadmium concentrate (stream 2350) is pelletized and may then beintroduced to a molten metal bath and made into ingots (stream 1500).

Simonkolleite/Zinc Oxychloride Precipitation:

The present invention first produces a high grade Simonkolleite and thenpurposefully converts it to a high purity zinc oxide. This isaccomplished by adding water (stream 540) to dilute the solutionchloride concentration sufficiently to begin to reverse the leachreactions. The result is the spontaneous and unexpected precipitation ofapproximately 35% of the zinc without the addition of a base. In thesecond phase of this step, a base is added to precipitate the bulk ofthe remaining zinc in solution. A cold solution temperature is preferredbecause it produces almost pure Simonkolleite, recovers approximately95% of the zinc in solution, and makes the determination of the amountof base to be added a direct function of the remaining zinc in solution.

The selection of the base is determined by the economics and the desiredend products of the entire process. Burnt lime (calcium oxide slurriedto make calcium hydroxide) is preferred because at this time, as it isless costly over all when considering the desired end products, it ispreferably used in the reduction furnace, and is a relativelyinexpensive raw source material. Alternatively, the inventive processlends itself to the use of sodium/potassium hydroxide because thesebases produce salt, which when extracted from the solution, become feedfor a chlor-alkali plant or other recycling process which can thenproduce sodium/potassium hydroxide and hydrochloric acid to be sold orused again in the process.

The Simonkolleite bearing solution can be separated, preferably with afilter, from the high chloride barren solution (stream 550) and may behot washed at 90° C. with, preferably, a calcium chloride solution(stream 810) to remove potassium chloride and to retain potassium insolution. The Simonkolleite is then water washed (stream 540) to removeresidual soluble chlorides to facilitate the quality of theSimonkolleite during the conversion to zinc oxide. The spent wash water(stream 550) is then recycled. Having the lowest possible level ofsoluble chlorides in the Simonkolleite/zinc oxychlorides after it isreslurried assures the lowest possible level of chlorides in the finalzinc oxide product when processed, according to the process disclosedherein.

H. Dilution of Filtrate Solution with water:

-   -   1. This step contains the solutions from two separate runs that        have completed the cementation step.    -   2. If the temperature of the solution is above room temperature,        turn on the cooling coil to cool the solution down to room        temperature (20-25° C.).        -   a) Remove a 50 ml sample (Label: ZnOx Liquid Before (before            dilution))            -   (1) Record the time and date.            -   (2) Measure the pH.            -   (3) Conduct an ICP analysis.            -   (4) Determine the solution volume and specific gravity                at room temperature.    -   3. Obtain a sample of the solution in Zinc Oxychloride Spent        Wash Tank 30 (recycled from prior runs) before using it for        dilution. (Label: ZnOx Dilution Water)        -   a) Record the time and date.        -   b) Measure the pH.        -   c) Conduct an ICP analysis.    -   4. Add water stored in Zinc Oxychloride Spent Wash Tank 30 to        Zinc Oxychloride Precipitation Tank 26 until the specific        gravity of the Zinc Oxychloride Precipitation Solution is 1.39        at room temperature (stream 560).    -   5. Mix for 60 minutes.    -   6. Determine the total solution volume at SG=1.39 and room        temperature.    -   7. Remove a 125 ml sample of solution (Label: ZnOx Liquid After        (after dilution))        -   a) Record the time and date.        -   b) Measure the pH.        -   c) Conduct an ICP analysis of the diluted solution.    -   8. Check with Lab Personnel to determine the level of Zn left in        solution.    -   9. The amount of lime required is calculated by the following        formula: Lime (18.9% calcium hydroxide)Solution (g)        Required=[Solution Vol. (L) at SG 1.39 X zinc (mg/kg) less        500(mg/kg)] X 0.0053214)

This step produces predominantly Simonkolleite (a zinc oxychloride) andsmall amounts of zinc oxide, depending on conditions, which is believedto occur according to the following reactions:5Ca⁺⁺+2ZnCl₃ ⁻+8Ca(OH)₂Zn₅Cl₂O₄.5H₂O+13CaCl₂  Equation 7Ca⁺⁺+2ZnCl₃ ⁻+2Ca(OH)₂2ZnO+3CaCl₂+2H₂  Equation 8ZnCl₂+Ca(OH)₂ZnO+CaCl₂+H₂O, 2  Equation 9wherein the Simonkolleite is Zn₅Cl₂O₄.5H₂₀.

I. Zinc Oxychloride Precipitation Step 2

-   -   1. Obtain a sample of the lime used in this step (Label: ZnOx        Lime)        -   a) Record the time and date.        -   b) Measure the pH.        -   c) Conduct an ICP analysis.    -   2. Measure and add the amount of Lime calculated in step G to        the Zinc Oxychloride Precipitation Tank 26 at room temperature        (stream 9100).        -   a) Record the total weight of Lime used and the time of            addition.    -   3. Mix the solution vigorously for one hour.    -   4. Remove a 125 ml sample of the slurry at 45 minutes and at 60        minutes after the lime addition (Label: ZnOx Slurry @ Time        ______).        -   a) Record the time and date the sample is taken.        -   b) Measure the pH.        -   c) Conduct an ICP analysis of the slurry.    -   5. Filter the solution through the Larox filter 28. The solution        may be filtered in several batches (stream 750). To filter the        ZnOx Slurry in 3 batches divide the total slurry volume in 3 and        do step 5-23 for each batch.        -   a) Attach the green hose to the outlet located at the bottom            of Zinc Oxychloride Precipitation Tank 26.        -   b) Place a clean filter zinc oxychloride/ZnO cloth into            filter 28. Make sure the cloth is centered before closing            the filter.        -   c) Push down on the end labeled “pump” on the hydraulic pump            located on the filter table. Allow the air to pump until the            pressure reading on the pressure gauge is close to the red            line. DO NOT pump past the red line.        -   d) Close the valve located behind the pressure gauge of the            hydraulic pump. This will help to prevent leaks through the            filter.        -   e) Make sure all valves on the filter are closed        -   f) Open the Zinc Oxychloride Precipitation Tank outlet            valve.        -   g) Turn on the pump.        -   h) Collect the filtrate in buckets.    -   6. Transfer the filtrate to Evaporation Tank 32 (stream 800 or        820).    -   7. Check the level of solution inside the tank and detach the        green hose when the tank is empty.    -   8. Allow the pump to continue pumping until the sound of the        pump starts to change.    -   9. Open the valve (yellow lever) to the Larox filter diaphragm        pump.    -   10. Open the valve (blue lever) located on the left side of the        pressure gauge on the Larox filter diaphragm pump. This allows        water to be pumped into the diaphragm and press the solids        inside the filter to “squeeze” more fluids out of the cake.    -   11. Slowly open the outlet valve and pump the remaining slurry        in the lines into a bucket.    -   12. Turn off the pump.    -   13. Dispose of the slurry properly.        -   a) Determine the total filtrate volume.        -   b) Obtain a 50 ml sample of filtrate (Label: ZnOx Filtrate).            -   (1) Record the time and date.            -   (2) Measure the pH.            -   (3) Conduct an ICP analysis.    -   14. Turn on the water pump to the water tower.    -   15. Close the valve (yellow lever) on the Larox filter diaphragm        pump.    -   16. Slow open the water valve to the filter and close the valve        (blue lever) on the Larox filter diaphragm pump.    -   17. Use 25 L (total) of H₂O to wash the Zinc Oxychloride Cake        (stream 540).        -   a) Close the filter outlet valve to stop the flow of            filtrate when 3 L of wash water has been collected.        -   b) Allow the solids to soak for two minutes.        -   c) Open the filter outlet valve.        -   d) Wash the cake with an additional 5.3 L (for each cake).    -   18. Open the valve (yellow lever) to the Larox filter diaphragm        pump.    -   19. Open the valve (blue lever) located on the left side of the        pressure gauge on the Larox filter diaphragm pump. This allows        water to be pumped into the diaphragm and press the solids        inside the filter to “squeeze” more fluids out of the cake.    -   20. Determine the total wash water volume. Theoretical volume=25        L total (the sum of wash water for 3 cakes).    -   21. Obtain a wash water sample (after washing the solids)        (Label: ZnOx Spent Wash Water).        -   a) Record the time and temperature of wash water solution.        -   b) Measure the pH.        -   c) Conduct an ICP analysis.    -   22. Close the valve (yellow lever) on the Larox filter diaphragm        pump.    -   23. Slowly open the air valve to the filter and close the valve        (blue lever) on the Larox filter diaphragm pump.    -   24. Allow the air to flow through filter for 1 minute.    -   25. Close the air valve to the filter.    -   26. Open the filter by pushing the end labeled “release” on the        hydraulic pump.    -   27. Carefully remove the filter cloth and the cake from the        filter (stream 1000).    -   28. Obtain a sample of cake (Label: ZnOx Cake).        -   a) Record the total weight of the cake.        -   b) Record the thickness of the cake.        -   c) Conduct an ICP analysis.    -   29. Store the solids inside a plastic bag. Label the bag        properly.

Evaporator Crystallization:

By the evaporation of water from the filtrate of the Zinc OxychloridePrecipitation the NaCl and KCl crystallize out when the correctconcentration of CaCl₂ to water is reached for the next leach. Bydropping the temperature below 90° C. the CaCl₂ concentration can becontrolled by crystallizing out excess CaCl₂ which builds up in thesystem due to the entry of chlorides in the feed. A temperature of 80°C. is preferred.

J. Recycle “Leach Salt Solution” Evaporator:

-   -   1. Transfer the solution in filter 28 into evaporator 32.    -   2. Record the volume and specific gravity of the solution        transferred.    -   3. Boil the sample inside the evaporator at 130° C. for 4 hours        to evaporate out the water.    -   4. Cool the remaining mixture down to 90° C. with the        evaporator's cooling system before filtering.    -   5. Check that the specific gravity of the solution inside        evaporator 32 is now 1.51, at a temperature of approximately 90°        C.    -   6. Cool the solution further to 80° C.    -   7. Remove 100 ml of sample (Label: Evap. Slurry).    -   8. Record the time and temperature.    -   9. Conduct an ICP analysis of the sample.    -   10. Filter the rest of the solution using the Larox Filter        (stream 850).

K. Pre-Heat Filter

-   -   1. Place a clean filter leach cake cloth into filter 16. Make        sure the cloth is centered before closing the filter.    -   2. Push down on the end labeled “pump” on the hydraulic pump        located on the filter table. Allow the air to pump until the        pressure reading on the pressure gauge is close to the red line.        DO NOT pump past the red line.    -   3. Close the valve located behind the pressure gauge of the        hydraulic pump. This will help to prevent leaks through the        filter.    -   4. Place the filter insulation (hereinafter the silver/gray        cover) over the filter and tighten the wires around the claps.    -   5. Place the heat gun at the bottom of the filter, pointing up        towards the filter. The heat gun, as well as all of the other        pre-heating steps, will help to prevent CaCl₂ precipitation in        the salt cake and to speed up the rate of filtration.    -   6. Turn on the water pump switch.    -   7. Slowly open the water valve and allow the water to flow        through the filter for one minute.    -   8. Close the water valve and open the air valve to blow out        excess water in the filter.    -   9. Close the air valve.    -   10. Attach the hose exiting reactor 20 to the green hose of the        filter pump.    -   11. Open the air bleed valve (hereinafter the black valve)        located to the left of the reactor pressure gauge to release any        vapor pressure that had built up inside reactor 20.    -   12. Unlatch the reactor door latches and insert the outlet hose        of the filter into the reactor. Secure the hose with plastic        cable latches.    -   13. Make sure all the valves on the filter are closed.    -   14. Open the reactor outlet valve located at the bottom of        reactor 20.    -   15. Turn on the pump to the filter.    -   16. Slowly open the valve to the outlet hose. This allows the        hot slurry to circulate from the pump line and back into the        reactor. The slurry will heat up the pump line as it circulates.    -   17. Allow the slurry to circulate in the lines for a couple of        minutes before proceeding to the Salt Slurry Filtration step.

L. Salt Slurry Filtration (NaCl/KCl Cake & Recycled CaCl₂ Solution).

-   -   1. Maintain a solution temperature of 80° C.    -   2. Close the outlet valve slightly to decrease the amount of        solution that flows back into the reactor.    -   3. Slowly open the valve to the filter and allow the slurry        pressure to reach a pressure of between 60-90 psi.    -   4. Collect the filtrate in buckets (stream 900)    -   5. Close the filter outlet valve.    -   6. Check the level of solution inside the reactor and allow the        pump to continue pumping until the sound of the pump starts to        change.    -   7. Close the reactor outlet valve.    -   8. Open the valve (yellow lever) to the Larox filter diaphragm        pump.    -   9. Open the valve (blue lever) located on the left side of the        pressure gauge on the Larox filter diaphragm pump. This allows        water to be pumped into the diaphragm and press the solids        inside the filter to “squeeze” more fluids out of the cake.    -   10. Slowly open the filter outlet valve and pump the remaining        slurry in the lines into a bucket.    -   11. Turn off the pump.    -   12. Detach the green hose.    -   13. Measure and record the volume, specific gravity and the        weight of the remaining slurry.    -   14. Dispose of the remaining slurry properly.    -   15. Close the valve (yellow lever) on the Larox filter diaphragm        pump.    -   16. Slow open the air valve to the filter and close the valve        (blue lever) on the Larox filter diaphragm pump.    -   17. Allow the air to flow through filter for one minute.    -   18. Close the air valve to the filter.    -   19. Open the filter by pushing the end labeled “release” on the        hydraulic pump.    -   20. Carefully remove the filter cloth and the cake from the        filter (stream 1700).    -   21. Obtain a sample of cake (Label: NaCl/KCl Cake).        -   a) Record the total weight of the cake.        -   b) Record the thickness of the cake.        -   c) Conduct an ICP analysis.    -   22. Store the solids inside a plastic bag. Label the bag        properly.    -   23. Determine the filtrate volume.    -   24. Determine the specific gravity of the filtrate.    -   25. Remove a 50 ml sample of the filtrate. (Label: Salt        Filtrate).        -   a) Record the time and reactor temperature.        -   b) Measure the pH, mV, and temperature of the sample in the            lab.        -   c) Conduct an ICP analysis.        -   d) Note any observations.    -   26. Check with the Lab personnel to ensure that the Ca level in        the filtrate is above 170,000 mg/kg before the filtrate is used        as leach solution for the next run.    -   27. Transfer the filtrate into the CaCl₂ tank or to storage        tank(s) and hold at a temperature of approximately 90° C.+.    -   28. Adjust the specific gravity of the solution to 1.49 by        adding water.

Zinc Oxychloride to Zinc Oxide Conversion:

The prior art in U.S. Pat. No. 1,863,700 teaches that zincoxide/oxychlorides produced from simple precipitation with calciumhydroxide contain as little as 1% chlorides and 3-5% calcium. The patentteaches that to reduce the calcium that hydrochloric acid must be addedto dissolve the calcium. The patent further teaches a process where azinc product can be produced with 1-2% calcium and 0.8% chlorides at100° C. Such a product containing 1% chlorides thus mathematicallycontains 7% Simonkolleite or other oxychlorides based on the chloridestoichiometry. Accordingly the product comprises less than 92% pure zincoxide, which is unacceptable for use as rubber grade zinc oxide.

The present invention starts with Simonkolleite/zinc oxychloride. TheSimonkolleite/zinc oxychloride and most zinc hydroxide present isthermally broken down between 125° C. and 170° C. The reaction goes tocompletion in the presence of a base as the zinc oxide precipitatesunder the elevated temperatures. We have learned that the solublechloride content of the solution in which the Simonkolleite isreslurried must be kept below a certain concentration through use ofdilution water so that the reaction with the base will minimize theoxychloride content in the final zinc oxide product. This is especiallytrue when using lime as the base because of the need to minimize excesscalcium in the final product. When using sodium or potassium hydroxide,a stoichiometric excess of base may be added (0-30% excess, with 10%being preferred) and thus less dilution water (50% preferred) isrequired. The final step is to dry the zinc oxide at a sufficienttemperature to decompose the remaining unreacted Simonkolleite, zinchydroxide and zinc hydrates and evaporate the remaining water at apreferred temperature of approximately 200° C. to decompose 1-4% zinchydrates and produce 99% plus pure zinc oxide.

M. ZnO Conversion:

-   -   1) Reslurry the zinc oxychloride cake collected from step H with        10 L of fresh H₂O.    -   2) Add 43 L of H₂O to reactor 60 and heat to 90° C. Fresh water        is used in the pilot plant process but in an actual plant, spent        ZnO wash water from the previous runs will be used as the        re-slurry water (stream 530).    -   3) Calculate the amount of lime required for the ZnO conversion        using the following formula:        Amount of Lime Required (kg)=Dry weight of Zinc Oxychloride        Cake(kg)*(0.3919).    -   4) Add the Zinc Oxychloride Re-slurry Slurry to reactor 60 at a        temperature of 90° C. (stream 1100).    -   5) The chemistry is as follows:        Zn₅Cl₂O₄.5H₂O+Ca(OH₂)5ZnO+CaCl₂₊₆H₂O  Equation 7

This reaction will only proceed to full completion when the chlorideconcentration is sufficiently low and sufficient time, temperature andmixing have occurred.

-   -   6) Obtain a sample of lime used in this step (Label: ZnO Lime).        -   a) Record the time and temperature of the solution.        -   b) Measure the pH.        -   c) Conduct an ICP analysis.    -   7) Measure and record the weight. Add the lime slurry to reactor        60 at a temperature of 90° C. (stream 9150).    -   8) Latch and tighten the bolts around the reactor door.    -   9) Record the time of addition and the temperature of reactor        60.    -   10) Allow the reactor to heat to 170° C.    -   11) Allow the reaction to run for an hour.    -   12) Filter the slurry with Larox filter 38 (stream 1150).        -   a) Open the black air bleed valve located to the left of the            reactor pressure gauge to release any vapor pressure that            has built up inside the reactor or cool the reactor to 100°            C.        -   b) Attach the green hose to the outlet located at the bottom            of the reactor.        -   c) Make sure all the valves on the filter are closed.        -   d) Open the reactor outlet valve.        -   e) Turn on the pump to the filter.    -   2. Collect the filtrate in buckets (stream 580)    -   3. Check the level of solution inside the tank and detach the        green hose when the tank is empty.

4. Allow the pump to continue pumping until the sound of the pump startsto change.

5. Open the valve (yellow lever) to the Larox filter diaphragm pump.

6. Open the valve (blue lever) located on the left side of the pressuregauge on the Larox filter diaphragm pump. This allows water to be pumpedinto the Larox filter diaphragm and press the solids inside the filterto “squeeze” more fluids out of the cake.

7. Slowly open the outlet valve and pump the remaining slurry in thelines into a bucket

-   -   8. Turn off the pump.    -   9. Dispose of the slurry properly.    -   10. Remove a 50 ml filtrate sample (Label: ZnO Filtrate).        -   a) Record the time and date the sample is taken.        -   b) Measure the pH, mV and temperature of the sample.        -   c) Conduct an ICP analysis.    -   11. Turn on the water pump to the water tower.    -   12. Close the valve (yellow lever) on the Larox filter diaphragm        pump.    -   13. Slow open the water valve to the filter and close the valve        (blue lever) on the Larox filter diaphragm pump.    -   14. Use 23.5 L (total) of H2O to wash the ZnO Cake (stream 510).        -   a) Close the filter outlet valve to stop the flow when 3 L            of wash water have been collected.        -   b) Allow the solids to soak for one minute.        -   c) Open the filter outlet valve.        -   d) Repeat steps a-c when 10 L of wash water has been            collected.        -   e) Repeat steps a-c when 15 L of wash water has been            collected.    -   15. Open the valve (yellow lever) to the Larox filter diaphragm        pump.    -   16. Open the valve (blue lever) located on the left side of the        pressure gauge on the Larox filter diaphragm pump. This allows        water to be pumped into the Larox filter diaphragm and press the        solids inside the filter to “squeeze” more fluids out of the        cake.    -   17. Determine the wash water volume. Theoretical volume=23.5 L        total.    -   18. Obtain a wash water sample (after washing the solids)        (Label: ZnO Cake Wash H₂O).        -   a) Record the time and temperature of the wash water            solution.        -   b) Measure the pH.        -   c) Conduct an ICP analysis.    -   19. Place the spent wash water into Spent ZnO Wash Tank (stream        520).    -   20. Close the valve (yellow lever) on the Larox filter diaphragm        pump.    -   21. Slow open the air valve to the filter and close the valve        (blue lever) on the Larox filter diaphragm pump.    -   22. Allow the air to flow through filter for one minute.    -   23. Close the air valve to filter.    -   24. Open the filter by pushing the end labeled “release” on the        hydraulic pump.    -   25. Carefully remove the filter cloth and the cake from the        filter (stream 1200).    -   26. Obtain a sample of cake (Label: ZnO Cake).        -   a) Record the total weight of the cake.        -   b) Record the thickness of the cake.        -   c) Conduct an ICP analysis.    -   27. Store the solids inside a plastic bag. Label the bag        properly.    -   28. Transfer the solid mixture to the ZnO oven and dry to        200° C. (stream 1300).    -   29. Analysis:        -   a) Conduct an ICP analysis of the dry ZnO mixture.        -   b) Analyze the mixture for Chlorides.

Iron/Steel Product Step:

Option 1 (FIG. 1): Preferred Method:

The calcium rich leach cake from the leach reactor filter 16 is directlyvaluable in iron/steel production due to its calcium content: or it ispreferably blended with the iron bearing material, i.e., EAF dust KO-61to enhance the removal of non-ferrous metals as the iron is reducedmaking it suitable for steel production feed stock

N. Form Briquette

-   -   1. Analyze the leach cake and the EAF dust.    -   2. Mix the EAF dust (100 grams) by adding pulverized coke (15.7        grams) or carbon dust, dried calcium rich leach cake (18.8        grams) and additional lime and magnesium with a small amount of        water (10 grams) based on the stoichiometric demand to, produce        a suitable “green” ball pellet with the structural strength        required, using standard industry practice for direct oxidation        reduction of the metal oxides.    -   3. Vary the chemistry to obtain direct reduced iron or Flux Cake        specifications of the steel mill customer and to maintain 95%        plus metallization.    -   4. Pelletize the green ball the material to a size suitable for        charge to the furnace to be used (¾ inched diameter or larger is        preferred).    -   5. Pre-heat the ball at 200-400° C., preferably 275° C., for        feed directly to a blast furnace or the preferred direct reduced        iron furnace.

O. Direct Reduced Iron Furnace

-   -   1. Process the Briquette/Pellet at 1000-1100° C., preferably        1050° C., in a reducing atmosphere for 15 to 75 minutes,        preferably 45 minutes, according to standard industry practice.        The preferred conditions create an iron/calcium material which        is over 95% metallized and has had 99% of the zinc and lead        removed.    -   2. The reduction furnace product is briquetted or pelletized        under pressure to produce a uniform solid over ¾ inch in        diameter; Cool in an oxygen-less atmosphere with non-contact        cooling (may be directly water cooled—this is not preferred as        it may fracture the briquette).    -   3. Analyze the resulting product.

P. Flue dust from the Briquetting or Direct Reduction furnace may berecycled to the hydrometallurgical process feed (preferred) or sold tothird parties for non-ferrous metal production.

Note: This metal oxide reduction step may be placed at the beginning ofthe process to produce a zinc rich furnace flue dust suitable as feedfor this process (FIGS. 3 and 4) (preferred). When so doing, the calciumrich leach cake produced from the hydrometallurgical portion of theprocess is fed into the “green” ball pellet feeder.

Referring now to FIG. 2, the process described in FIG. 1 is shown exceptthat caustic is used, instead of lime, for feed to zinc oxide reactor60. Additionally, a caustic regeneration loop is provided which producescaustic and hydrochloric acid for both inside usage and outside sale. Inparticular, tank 62 contains caustic. Filter 40 feeds calcium chloridein stream 900 to leach reactor 20, and feeds sodium chloride andpotassium chloride in stream 1700 to a salt re-slurry tank 80.Evaporator 32 feeds water in a stream 512 to salt re-slurry tank 80.Tank 80 dissolves the salts and feeds the dissolved salt solution in astream 1710 to a calcium ion removal system 82. An ion exchange isconducted in tank 82 to remove the calcium ions from solution. Tank 82then feeds the sodium/potassium solution without the calcium ions in astream 1720 to a chlor-alkali plant 84. Plant 84 produces caustic whichis fed in stream 1730 to caustic tank 62 and produces hydrochloric acidwhich is fed in a stream 1740 to a hydrochloric tank 86 for outsidesale. A portion of the hydrochloric acid may be used to maintainchloride balance in stream 200 or to feed stream 9200 to cadmium leachtank 42. During this process, caustic is reacted with stream 700 in tank26 to make Simonkolleite and in reactor 60, the Simonkolleite is reactedwith the caustic to produce zinc oxide, sodium chloride and water.Accordingly, reactor 26 and reactor 60 produce sodium/potassiumchloride, instead of calcium chloride, as in FIG. 1. In this process,calcium chloride at stream 1800 typically is not produced.

Referring to FIG. 3, the process described in FIG. 1 is shown exceptthat the sequence of hearth furnace 14 and electric arc furnace 12 isreversed. Accordingly, hearth furnace 14 feeds metallized ironbriquettes in stream 1900 to EAF 12. EAF 12 feeds EAF Red Dust in stream9900 to dust storage tank 18. K061 is processed in leach reactor 20 andthe dregs are dried and sold as magnetite or used to make briquettes instream 1600 for feed to the reduction hearth furnace 14. Again thepresence of calcium hydroxide from the leach will enhance non-ferrousmetal recover and metallization of the iron for an improved product1900. Any non-ferrous dust generated in hearth furnace 14 is added totank 18. The remainder of the process is the same as that shown in FIG.1.

Referring to FIG. 4, the process described in FIG. 2 is shown exceptthat the sequence of hearth furnace 14 and electric arc furnace 12 isreversed. Accordingly, hearth furnace 14 feeds metallized ironbriquettes in stream 1900 to EAF 12. EAF 12 feeds EAF Red Dust in stream9900 to dust storage tank 18. K061 is processed in leach reactor 20 andthe dregs dried and sold as magnetite or are used to make briquettes instream 1600 for feed to hearth furnace 14. Because calcium is notgenerated by the leach process, to obtain the benefits of calciumaddition it must be separately added (not shown). The remainder of theprocess is the same as that shown in FIG. 2.

FIG. 5 shows a detailed schematic flow diagram of the Simonkolleiteproduction and zinc oxide production portions of the overall process ofFIG. 1.

FIG. 6 shows a detailed schematic flow diagram of the zinc oxide fromzinc oxychloride portion of the overall process of FIG. 1.

FIG. 7 shows a graph which compares the zinc oxide produced by thepresent inventive process with zinc oxide prepared by a standardprocess. According to the graph, the zinc oxide produced by the presentinvention is an exact compositional match to the zinc oxide produced bythe French Process prior art process.

Table 1 shows tests of rubber compounds using zinc oxide produced by theinventive process.

Table 2 shows tests of rubber compounds using zinc oxide produced by theinventive process.

Table 3 shows tests of rubber compounds using zinc oxide produced by theinventive process.

Table 4 shows tests of rubber compounds using zinc oxide produced by theinventive process.

Table 5 shows tests of rubber compounds using zinc oxide produced by theinventive process.

Table 6 shows an elemental chemical analysis of the components of thehearth dust in stream 9950 for several different runs.

Table 7 shows an elemental chemical analysis of the components of thecalcium chloride in streams 200 (upper sub-table) and 210 (lowersub-table) for several different runs.

Table 8 shows an elemental chemical analysis of the components of theleach cake in stream 1600 for several different runs.

Table 9 shows an elemental chemical analysis of the components of theheavy cementation in stream 2000 for several different runs.

Table 10 shows an elemental chemical analysis of the components of thezinc chloride complex solution in stream 700 for several different runs.

Table 11 shows an elemental chemical analysis of the components of thelime in stream 9100 for several different runs.

Table 12 shows an elemental chemical analysis of the components of thefilter cake material pressed in filter 38 before washing with stream510, for several different runs.

Table 13 shows an elemental chemical analysis of the components ofstream 800 from filter 28 for several different runs.

Table 14 shows an elemental chemical analysis of the components ofstream 1700 from filter 40 for several different runs.

Table 15 shows an elemental chemical analysis of the components ofstream 580 from filter 38 for several different runs.

Table 16 shows an elemental chemical analysis of the components ofstream 1200, the final process product, for several different runs.

Table 17 shows the iron briquetted product analysis for various testconditions vs. that for the KO-61 feed without calcium additionconducted in Round 1 testing. It can be observed that the zinc and leadlevels are far lower when calcium is added then when it is not added tothe reduction furnace feed.

Table 18 shows the iron/calcium briquette analysis from Round 2 testing,under more optimized conditions where the amount of zinc and lead havebeen reduced by over 99%.

While the preferred embodiment of the invention has been illustrated anddescribed, it will be appreciated that various changes can be madetherein without departing from the spirit and scope of the invention.

TABLE 1 ODR ODR ODR ODR ODR MS Tensile Tensile Tensile Tmin Tmax ts1 t90CRI 15 Hardness stress @ 300% stress @ break strain @ break NR/ZnOCOMPOUND STUDY - AMR NR-CONTROL 7.1 34.7 3.26 7.24 25.12 23.42 57.9 14904010 539% NR-25/150 6.7 34.5 3.16 7.08 25.51 22.00 56.7 1502 4000 535%NR-25/150 LOW 6.7 34.0 3.21 7.03 26.18 21.58 57.0 1478 4130 554%NR-25/150 HIGH 6.9 35.6 3.08 7.02 25.38 22.33 59.7 1631 3860 514%NR-TEST1 Stearic acid 6.8 34.6 3.22 7.54 23.14 22.42 57.4 1536 4130 553%NR-TEST2 Poly 6.6 34.2 3.21 7.31 24.39 22.42 57.1 1531 4220 556%NR-TEST3 TEA 6.8 34.7 3.11 7.01 25.64 20.83 56.3 1437 3930 535% NR-TEST4No additive 6.5 34.5 3.19 7.15 25.26 22.08 57.8 1479 4010 539% NR-w/oZnO 6.4 16.2 3.23 11.77 11.71 SBR/ZnO COMPOUND STUDY - AMR SBR-CONTROL10.4 44.2 4.23 16.75 7.99 37.83 71.7 2359 3540 411% SBR-25/150 10.7 44.74.13 16.85 7.86 37.00 73.0 2304 3330 396% SBR-25/150 LOW 10.4 43.5 4.2416.53 8.14 35.17 70.6 2350 3440 401% SBR-25/150 HIGH 10.7 45.2 4.2416.37 8.24 39.17 73.0 2448 3530 401% SBR-TEST1 10.4 44.4 4.02 16.54 7.9937.25 71.8 2427 3490 399% SBR-TEST2 10.4 43.8 3.83 15.84 8.33 37.75 72.42390 3430 395% SBR-TEST3 10.3 44.6 4.02 16.26 8.17 38.17 70.2 2414 3580406% SBR-TEST4 10.3 44.1 4.12 16.64 7.99 37.25 70.6 2368 3710 427%SBR-w/o ZnO 10.6 33.0 4.20 23.99 5.05 CR/ZnO COMPOUND STUDY - AMRCR-CONTROL 6.1 44.5 1.23 3.50 44.05 4.92 67.9 1650 3690 578% CR-25/1505.8 44.8 1.31 3.89 38.76 5.67 67.6 1684 3860 586% CR-25/150 LOW 5.9 43.81.42 3.81 41.84 6.50 65.4 1613 3780 582% CR-25/150 HIGH 6.7 48.7 1.325.89 21.89 5.00 67.6 1825 3440 524% CR-TEST1 6.1 44.0 1.32 3.87 39.235.67 67.0 1668 3550 591% CR-TEST2 6.0 44.2 1.36 3.88 40.00 6.50 67.51566 3760 592% CR-TEST3 6.1 44.3 1.31 4.18 34.84 5.33 67.6 1677 3760582% CR-TEST4 6.0 44.7 1.31 3.81 40.00 5.58 67.9 1709 3790 583% CR-w/oZnO 5.5 37.4 3.21 6.82 27.70

TABLE 2 REPLACEMENT TREAD COMPOUND - SYNTHETIC 3-1 3-2 3-3 3-4 3-5 3-63-7 3-8 3-9 3-10 SBR 1712 96.0 96.0 96.0 96.0 96.0 96.0 96.0 96.0 96.098.0 BUDENE 1207 30.0 30.0 30.0 30.0 30.0 30.0 30.0 30.0 30.0 30.0CARBON BLACK 65.0 65.0 65.0 65.0 65.0 65.0 65.0 65.0 65.0 65.0 KADOX 9111.0 4.0 8.0 AMR-ZNO 1.0 4.0 8.0 ACT-ZNO-A 1.0 8.0 ACT-ZNO-B 1.0 8.0STEARIC ACID 2.0 2.0 2.0 2.0 2.0 2.0 2.0 2.0 2.0 2.0 AROMATIC OIL 5.05.0 5.0 5.0 5.0 5.0 5.0 5.0 5.0 5.0 SPPD 1.5 1.5 1.5 1.5 1.5 1.5 1.5 1.51.5 1.5 MC WAX 1.0 1.0 1.0 1.0 1.0 1.0 1.0 1.0 1.0 1.0 TMTD 0.2 0.2 0.20.2 0.2 0.2 0.2 0.2 0.2 0.2 OBTS 1.1 1.1 1.1 1.1 1.1 1.1 1.1 1.1 1.1 1.1SULFUR 1.6 1.6 1.6 1.6 1.6 1.6 1.6 1.6 1.6 1.6 HEAVY DUTY COMPOUND -NATURAL 4-1 4-2 4-3 4-4 4-5 4-6 4-7 4-8 4-9 4-10 SMR-10 (NR) 100.0 100.0100.0 100.0 100.0 100.0 100.0 100.0 100.0 100.0 SILICA 15.0 15.0 15.015.0 15.0 15.0 15.0 15.0 15.0 15.0 CARBON BLACK 40.0 40.0 40.0 40.0 40.040.0 40.0 40.0 40.0 40.0 KADOX 911 1.5 5.0 10.0 AMR-ZNO 1.5 5.0 10.0ACT-ZNO-A 1.5 10.0 ACT-ZNO-B 1.5 10.0 STEARIC ACID 1.0 1.0 1.0 1.0 1.01.0 1.0 1.0 1.0 1.0 TMQ 2.0 2.0 2.0 2.0 2.0 2.0 2.0 2.0 2.0 2.0 SPPD 3.03.0 3.0 3.0 3.0 3.0 3.0 3.0 3.0 3.0 MC WAX 2.0 2.0 2.0 2.0 2.0 2.0 2.02.0 2.0 2.0 AROMATIC OIL 5.0 5.0 5.0 5.0 5.0 5.0 5.0 5.0 5.0 5.0 SULFUR0.7 0.7 0.7 0.7 0.7 0.7 0.7 0.7 0.7 0.7 SULFASAN R 2.0 2.0 2.0 2.0 2.02.0 2.0 2.0 2.0 2.0 OBTS 1.1 1.1 1.1 1.1 1.1 1.1 1.1 1.1 1.1 1.1AIRCRAFT TYPE TREAD COMPOUND 5-1 5-2 5-3 5-4 5-5 5-6 5-7 5-8 5-9 5-10SMR-10 (NR) 100 100 100 100 100 100 100 100 100 100 CARBON BLACK 50 5050 50 50 50 50 50 50 50 KADOX 911 1.5 10 20 ARM-ZNO 1.5 10 20 ACT-ZNO-A1.5 20 ACT-ZNO-B 1.5 20 STEARIC ACID 3 3 3 3 3 3 3 3 3 3 TMQ 4 4 4 4 4 44 4 4 4 SPPD 2 2 2 2 2 2 2 2 2 2 PINE TAR 3.5 3.5 3.5 3.5 3.5 3.5 3.53.5 3.5 3.5 SULFUR 2.1 2.1 2.1 2.1 2.1 2.1 2.1 2.1 2.1 2.1 SULFASAN R0.5 0.5 0.5 0.5 0.5 0.5 0.5 0.5 0.5 0.5 OBTS 1 1 1 1 1 1 1 1 1 1

TABLE 3 TEST REPORT SUMMARY REPLACEMENT TREAD COMPOUND 3-1 3-2 3-3 3-43-5 3-6 3-7 3-8 3-9 3-10 ODR @ 320 F. Tm (in-lbf) 33.0 35.9 36.4 32.134.8 35.5 30.9 34.9 27.6 34.4 ts1 (min.) 5.27 6.11 5.99 5.19 6.11 5.885.18 5.01 4.19 5.36 tc90 (min.) 10.57 13.49 13.43 10.81 13.47 13.3310.66 12.34 8.14 11.66 CRI 18.52 13.55 13.44 17.80 13.59 13.42 18.2513.64 25.32 15.87 REVERSION (in-lbf/min.) 0 0 0 0 0 0 0 0 0 0 ODR @ 350F. Tm 29.3 32.2 33.0 28.5 31.2 33.7 30.1 32.9 26.3 32.2 ts1 2.7 2.7 2.62.5 2.8 2.7 2.6 2.4 2.3 2.6 tc90 4.9 5.7 5.6 4.8 5.7 5.6 5.1 5.4 4.2 5.3CRI 45.46 33.44 33.33 43.30 34.48 34.48 40.00 33.23 52.65 37.17REVERSION 0.06 0.04 0.04 0.05 0.04 0.04 0.06 0.04 0.08 0.04 MOONEYSCORCH t5 (min.) @ 257 F. 30.63 50.25 49.25 29.92 46.58 46.68 28.5047.08 23.83 39.58 TENSILE, ORIG, T95 @ 320 F. TENSILE STRESS (psi) 29402830 2840 3010 2860 2710 2980 2830 2790 2890 TENSILE @ 300% (psi) 11571282 1396 1138 1293 1395 1016 1328 775 1265 TENSILE STRAIN % 582% 525%505% 608% 546% 500% 642% 505% 859% 552% HARDNESS, Shore A 59.1 61.5 63.260.6 62.8 64.7 58 61.8 56.4 83.3 TENSILE, AGED 158 HRS @ 100 C., T95 @320 F. TENSILE STRESS 2300 2280 2250 2390 3100 2200 2350 2160 2320 2270TENSILE @ 300% TENSILE STRAIN 272% 260% 247% 283% 245% 242% 296% 233%278% 251% HARDNESS 72.5 71.8 73.2 70.5 74.6 72.6 72.6 73.8 70.6 75.5TENSILE, ORIG, 2X T95 @ 320 F. TENSILE STRESS 3100 2800 2970 3110 29502960 3050 2870 2850 2980 TENSILE @ 300% 1076 1214 1427 1064 1311 1360974 1411 763 1248 TENSILE STRAIN 618% 532% 518% 635% 538% 533% 654% 507%601% 568% HARDNESS 59.3 62.1 62.9 59.9 62.4 62.5 58.0 60.4 56.0 62.0TENSILE, AGED 188 HRS @ 100 C., 2X T95 @ 320 F. TENSILE STRESS 2460 23702410 2510 2360 2460 2490 2420 2440 2440 TENSILE @ 300% 2365 2327 22972353 2244 2370 2330 TENSILE STRAIN 312% 303% 295% 324% 302% 288% 329%297% 306% 314% HARDNESS 71.0 72.1 73.5 71.5 73.4 73.4 71.5 73.2 71.072.5 FLEXOMETER DELTA T, deg F. 107 83 97 110 95 98 126 83 80 97 BLOWOUTTIME E′ E″ TAN DELTA

TABLE 4 TEST REPORT SUMMARY OTR TREAD 4-1 4-2 4-3 4-4 4-5 4-6 4-7 4-84-9 4-10 ODR @ 320 F. Tm (in-lbf) 42.9 44.2 43.7 43.7 43.5 43.2 42.845.5 41.9 45.8 ts1 (min.) 9.56 9.25 8.37 9.32 9.21 9.31 8.68 8.29 8.078.11 tc90 (min.) 18.81 18.57 17.22 19.42 19.09 19.12 17.58 17.29 14.9815.99 CRI 10.81 10.73 11.30 9.00 10.12 10.20 11.24 11.11 14.48 12.69REVERSION (in-lbf/min.) 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.300.00 ODR @ 350 F. Tm 38.5 38.9 39.5 39.5 39.5 40.0 39.1 41.2 37.1 41.3ts1 2.1 2.2 2.1 2.2 2.1 2.1 2.2 2.1 2.0 2 tc90 4.1 4.2 3.9 4.3 4.2 4.24.2 4.1 3.6 3.9 CRI 50.00 50.02 55.25 47.62 47.41 47.41 50.02 50.0062.49 52.65 REVERSION 0.24 0.19 0.24 0.19 0.23 0.21 0.26 0.22 0.33 0.20MOONEY SCORCH t5 (min.) @ 280 F. 17.67 16.67 15.42 15.92 16.83 17.0015.50 15.08 13.75 13.75 TENSILE, ORIG, T95 @ 300 F. TENSILE STRESS (psi)3709 3480 3645 3572 3683 3603 3477 3531 3403 3356 TENSILE @ 300% (psi)1969 2087 2148 2057 2125 2172 2073 2158 1837 2173 TENSILE STRAIN % 499%451% 466% 481% 478% 457% 457% 458% 482% 432% HARDNESS, Shore A 67.5 69.769.8 69.5 70.3 69.3 69.0 70.5 68.6 71.3 TENSILE, AGED 168 HRS @ 100 C.,T95 @ 300 F. TENSILE STRESS 2320 2340 2430 2360 2300 2400 2110 2510 21702270 TENSILE @ 300% 2054 2147 2183 2092 2159 2150 2081 2194 2021 2148TENSILE STRAIN 340% 329% 335% 341% 321% 336% 306% 342% 324% 317%HARDNESS 74.3 75.3 75.3 76.3 75.6 74.2 75.4 77.6 74.2 75.3 TENSILE,ORIG, 2X T95 @ 300 F. TENSILE STRESS 3499 3352 3320 3244 3341 3209 33243279 2963 3291 TENSILE @ 300% 1820 1998 1971 2019 1914 2035 2038 22781648 2317 TENSILE STRAIN 496% 452% 449% 439% 464% 431% 440% 406% 457%400% HARDNESS 66.4 68.3 67.3 68.3 66.2 66.8 68.4 70.7 63.8 66.3 TENSILE,AGED 168 HRS @ 100 C., 2X T95 @ 300 F. TENSILE STRESS 2280 2370 21902130 2200 2270 2230 2200 1640 2310 TENSILE @ 300% 2018 2076 2101 19822071 2054 2031 2108 2161 TENSILE STRAIN 338% 343% 315% 323% 317% 332%330% 313% 266% 322% HARDNESS 71.9 73.9 73.6 74.3 74.6 75.0 73.0 75.472.3 75.2 FLEXOMETER DELTA T, deg F., 55/22.5/30 60 63 70 72 74 67 65 6179 72 BLOWOUT TIME E′ E″ TAN DELTA

TABLE 5 TEST REPORT SUMMARY AIRCRAFT TREAD 5-1 5-2 5-3 5-4 5-5 5-6 5-75-8 5-9 5-10 ODR @ 300 F. Tm (in-lbf) 38.1 42.1 42.2 38.1 40.2 41.8 38.442.3 32.4 42.3 ts1 (min.) 5.80 5.73 5.60 5.71 5.48 5.23 5.45 4.71 4.905.01 tc90 (min.) 12.03 13.57 14.31 12.34 13.19 12.73 11.85 13.34 7.9811.96 CRI 16.05 12.75 11.48 15.08 12.97 13.33 15.63 11.59 32.47 14.39REVERSION (in-lbf/min.) 0.29 0.00 0.00 0.20 0.00 0.00 0.30 0.00 0.410.00 ODR @ 350 F. Tm 38.3 38.1 38.0 34.6 36.9 37.9 35.4 39.3 30.4 38.2ts1 1.6 1.5 1.4 1.5 1.4 1.5 1.5 1.4 1.5 1.5 tc90 3.2 3.3 3.3 3.1 3.2 3.13.1 3.4 2.5 3.0 CRI 62.14 55.85 52.90 62.49 55.57 62.53 62.49 50.00100.04 66.24 REVERSION MOONEY SCORCH t5 (min.) @ 257 F. 25.75 24.8324.17 25.17 24.08 23.58 25.25 25.42 21.42 25.92 TENSILE, ORIG, T95 @ 300F. TENSILE STRESS (psi) 3870 3660 3580 3720 3670 3470 3620 3540 35503370 TENSILE @ 300% (psi) 1920 2230 2292 1920 2073 2210 1892 2219 13912161 TENSILE STRAIN % 515% 462% 441% 510% 486% 450% 501% 450% 578% 445%HARDNESS, Shore A 67.6 70.4 70.2 69.8 68.4 72.8 69.0 72.8 61.6 71.0TENSILE, AGED 168 HRS @ 100 C., T95 @ 300 F. TENSILE STRESS 2020 16801810 2100 1830 1870 1860 1810 1550 1690 TENSILE @ 300% TENSILE STRAIN225% 185% 199% 247% 224% 205% 206% 200% 193% 195% HARDNESS 76.7 77.977.1 75.6 77.0 76.8 76.7 77.8 73.8 79.1 TENSILE, ORIG, 2X T95 @ 300 F.TENSILE STRESS 3630 3450 3390 3630 3410 3360 3470 3420 3280 3240 TENSILE@ 300% 1656 2333 2277 1776 2160 2234 1638 2279 1153 2109 TENSILE STRAIN538% 420% 415% 519% 450% 432% 519% 426% 611% 435% HARDNESS 64.0 71.072.0 57.0 71.0 71.0 64.0 72.0 60.0 71.0 TENSILE, AGED 168 HRS @ 100 C.,2X T95 @ 300 F. TENSILE STRESS 1670 1910 1870 1580 1920 1900 1460 17201280 1710 TENSILE @ 300% TENSILE STRAIN 180% 191% 194% 180% 217% 204%161% 174% 143% 187% HARDNESS 72.8 76.9 77.0 74.6 77.4 78.3 74.3 77.474.0 79.6 FLEXOMETER DELTA T, deg F. 23 26 25 26 30 29 25 31 85 29BLOWOUT TIME E′ E″ TAN DELTA

TABLE 6 American Metals Recovery Sample AHD2 #1 AHD2 #2 AHD2 #3 AHD2 #4AVG AVG Dilution 0.5051/1000 0.5010/1000 0.5151/1000 0.5082/1000 #1-3#1-4 Sc 361.383 4.87 4.65 4.73 4.67 Ag 338.289 1183 1235 1135 1249 11841200 Al 308.215 212 206 187 221 202 207 As 188.979 33.00 23.9 18.8 36.3<10 <10 B 249.772 78.51 80.91 74.87 77.10 78.09 77.85 Ba 230.425 14.3714.10 14.02 15.30 14.16 14.45 Be 313.107 <10 <10 <10 <10 <10 <10 Ca317.933 1259 788 802 967 950 954 Cd 214.440 2463 2384 2294 2603 23742431 Co 228.616 <10 <10 <10 <10 <10 <10 Cr 283.563 68.81 67.49 65.3073.78 67.20 68.85 Cu 324.752 620.9 597 597 858 605 618 Fe 238.204 86538249 8265 9343 8389 8628 K 766.490 16553 15967 15689 17384 16070 16398Mg 279.077 692 678 684 753 685 702 Mn 257.610 850 821 815 900 829 847 Na330.237 18920 16941 16575 19394 16812 17458 Ni 221.648 <10 <10 <10 <10<10 <10 P 178.221 64.0 26.9 46.33 36.38 45.75 43.40 Pb 220.353 6915766813 66386 73297 67445 88908 S 180.669 5367 5416 5182 5619 5322 5396 Sb206.838 60.5 63.22 57.72 71.20 60.47 63.15 Se 196.028 <10 <10 <10 <10<10 <10 SI 251.611 346 356 433 624 378 440 TI 190.801 <10 <10 <10 <10<10 <10 V 292.402 <10 <10 <10 <10 <10 <10 Zn 208.200 599946 579276573631 635923 584284 597194

TABLE 7 American Metals Recovery Corp. Leach Salt Solution and LeachSalt Wash Run #1 Run #2 Run #3 Run #4 Run #10 Run #11 Run #12 Run #13Leach Salt Soln Date 071399 071399 071399 071399 071399 071399 071399I.D. R1&2LSR R3LSR R4LSR R10LSR R11LSR R11LSR R12&13LSR Sc 361.383 4.894.95 4.89 4.93 4.92 4.92 4.85 Ag 338.289 3.04 2.57 2.24 2.47 <1 <1 2.39Al 308.215 <1 <5 <5 <1 <1 <1 <1 As 188.979 <1 <5 <5 <1 <1 <1 <1 B249.772 70.36 66.80 64.38 68.92 44.67 44.67 70.76 Ba 230.425 <1 <1 <1 <1<1 <1 <1 Be 313.107 <1 <1 <1 <1 <1 <1 <1 Ca 317.933 177887 170739 174342170644 123165 123165 179271 Cd 214.440 <1 <1 <1 <1 <1 <1 <1 Co 228.616<1 <1 <1 <1 <1 <1 <1 Cr 283.563 <1 <1 <1 <1 <1 <1 <1 Cu 324.752 <1 <1 <1<1 <1 <1 <1 Fe 238.204 <1 1.27 <1 <1 <1 <1 <1 K 766.490 8179 7645 78468086 5659 5859 9112 Mg 279.077 19.21 12.50 19.19 16.45 9.26 9.26 <1 Mn257.610 <1 <1 <1 <1 <1 <1 <1 Na 330.237 3133 1162 2910 3066 2036 20362515 Ni 221.648 <1 <1 <1 <1 <1 <1 <1 P 213.617 <1 <2 <1 <1 <1 <1 <1 Pb220.353 <1 <1 <1 <2 3.60 3.60 <1 S 180.669 Sb 206.836 <1 <1 <1 <1 <1 <1<1 Se 196.026 <1 <1 <1 <1 <1 <1 <1 Si 251.611 <1 <1 <1 <1 8.41 8.41 5.67TI 190.801 <1 <1 <1 <1 <1 <1 <1 V 292.402 <1 <1 <1 <1 <1 <1 <1 Zn206.200 126.37 29.3 20.0 124.61 499.5 499.5 1170.1 1:10, Ca, K, Zn from1:100 Leach Salt Wash Date 071399 071399 071399 071399 071399 071399071399 I.D. R1&2LSR R1&2LSR R4LSR R10LSR R11LSR R12&13LSR R12&13LSR Sc361.383 4.89 4.89 4.89 4.93 4.92 4.85 4.85 Ag 338.289 3.04 3.04 2.242.24 <1 2.39 2.39 Al 308.215 <1 <1 <5 <1 <1 <1 <1 As 188.979 <1 <1 <5 <1<1 <1 <1 B 249.772 70.36 70.36 64.38 68.92 44.67 70.76 70.76 Ba 230.425<1 <1 <1 <1 <1 <1 <1 Be 313.107 <1 <1 <1 <1 <1 <1 <1 Ca 317.933 177887177887 174342 170644 123165 179271 179271 Cd 214.440 <1 <1 <1 <1 <1 <1<1 Co 228.616 <1 <1 <1 <1 <1 <1 <1 Cr 283.563 <1 <1 <1 <1 <1 <1 <1 Cu324.752 <1 <1 <1 <1 <1 <1 <1 Fe 238.204 <1 <1 <1 <1 <1 <1 <1 K 766.4908179 8179 7846 8086 5659 9112 9112 Mg 279.077 19.21 19.21 19.19 16.459.26 <1 <1 Mn 257.610 <1 <1 <1 <1 <1 <1 <1 Na 330.237 3133 3133 29103066 2036 2515 2515 Ni 221.648 <1 <1 <1 <1 <1 <1 <1 P 213.617 <1 <1 <1<1 <1 <1 <1 Pb 220.353 <1 <1 <1 <2 3.60 <1 <1 S 180.669 Sb 206.836 <1 <1<1 <1 <1 <1 <1 Se 196.026 <1 <1 <1 <1 <1 <1 <1 Si 251.611 <1 <1 <1 <18.41 5.87 5.87 TI 190.801 <1 <1 <1 <1 <1 <1 <1 V 292.402 <1 <1 <1 <1 <1<1 <1 Zn 206.200 126.37 126.37 20.0 124.61 499.5 1170.1 1170.1

TABLE 8 Leach Cake Run #1 Run #2 Run #3 Run #4 Run #6 Run #6 Run #7 Run#8 Date 070699R 070699R 070699R 070699R 070699R 070699R 070699R 070699RI.D. R1LC R2LC R3LC R4LC R5LC R6LC R7LC R8LC Sc 361.383 4.94 4.99 4.954.96 5.01 5.04 4.98 5.00 Ag 338.289 <15 <15 127.94 <15 <15 <15 <15 <15Al 308.215 306.68 247.71 142.37 202.83 124.98 364.51 200.21 255.49 As188.979 <15 <15 <15 <15 <15 <15 30.36 <15 B 249.772 <15 <15 <15 <15 <15273.31 266.08 157.38 Ba 230.425 <15 <15 <15 <15 <15 <15 <15 <15 Be313.107 <15 <15 <15 <15 <15 <15 <15 <15 Ca 317.933 440550 386712 407398398601 429743 396474 389290 442256 Cd 214.440 624.22 555.06 219.99462.98 159.45 475.49 198.36 623.61 Co 228.616 <15 <15 <15 <15 <15 <15<15 <15 Cr 283.563 102.34 79.96 <50 <50 31.59 110.88 56.14 80.73 Cu324.752 199.11 172.79 99.41 117.32 67.82 232.54 76.78 141.68 Fe 238.20413853 11686 6751 9973 5783 16968 8502 11910 K 766.490 <50 265.09 1337216.91 1399 306.40 <15 <15 M 279.077 1171 1339 965 1241 656.18 2122646.51 914 Mn 257.610 1253 1031 596.12 877 497.86 1454 743.21 1035 Na330.237 <50 <50 <50 <50 <15 <50 <50 <50 Ni 221.648 <50 <50 <50 <50 <50<50 <50 <50 P 213.617 <50 <50 <50 <50 <75 <50 82.86 <50 Pb 220.353 639500 465.45 424 688.47 701.02 621.91 630.76 S 180.669 Sb 206.836 <50 <50<50 <50 <50 <50 <50 <50 Se 196.026 <50 <50 <50 <50 <50 <50 <50 <50 Si251.611 959.64 717.25 347.27 527 <50 <50 1178.21 1193 TI 190.801 <50 <50<50 <50 <50 <50 <50 <50 V 292.402 <50 <50 <50 <50 <50 <50 <50 <50 Zn206.200 75414 49870 17428 31931 19633 63319 29289 42874 Leach Cake Run#9 Run #10 Run #11 Run #12 Run #13 Date 070699R 070699R 070699R 070699R070699R I.D. R9LC R10LC R11LC R12LC R13LC Sc 361.383 5.05 5.01 5.01 4.985.01 Ag 338.289 <15 <15 <15 <15 <15 Al 308.215 391.32 311.14 186.90145.07 221.85 As 188.979 <15 <15 <15 <15 <15 B 249.772 312.58 <50 235.73137.68 153.30 Ba 230.425 <15 <15 <15 <15 <15 Be 313.107 <15 <15 <15 <15<15 Ca 317.933 452411 433313 420944 401001 424539 Cd 214.440 376.73277.23 326.25 152.36 239.52 Co 228.616 <15 <15 <15 <15 <15 Cr 283.563121.49 <50 <50 <50 67.56 Cu 324.752 257.23 110.32 123.68 97.56 162.21 Fe238.204 17407 8447 8281 6266 9670 K 766.490 <15 1457 737.13 1058.43741.77 M 279.077 1369 1422 1048 493.15 749.50 Mn 257.610 1502 723.41718.56 544.34 848.42 Na 330.237 <50 <50 <50 <50 <50 Ni 221.648 <50 <50<50 <50 <50 P 213.617 <50 <50 <50 <50 <50 Pb 220.353 558.28 1058 663.15698.87 821 S 180.669 Sb 206.836 <50 <50 <50 <50 <50 Se 196.026 <50 <50<50 <50 <50 Si 251.611 1565 454.25 1082 686.80 832.92 TI 190.801 <50 <50<50 <50 <50 V 292.402 <50 <50 <50 <50 <50 Zn 206.200 75343 25832 3042032960 51817

TABLE 9 American Metals Recovery Corp. Bulk Cement Stage 1 Solids BC Stg1 solids (uw) Run #1 Run #2 Run #3 Run #4 Run #5 Run #7 Run #8 Run #9Date 070799R 070799R 070799R 070799R 070799R 070799R 070799R 070799RI.D. R1BCS1C R2BCS1C R3BCS1C R4BCS1C R5BCS1C R7BCS1C R8BCS1C R9BCS1C Sc361.383 5.12 5.21 5.11 5.16 5.18 5.04 5.07 5.00 Ag 338.289 20.91 19.1619.48 26.19 19.59 29.87 52.21 18.47 Al 308.215 96.78 <50 <50 92.2 176.7130.3 233.1 137.89 As 188.979 <10 <50 <10 <10 <10 <10 <10 <10 B 249.772<50 113 <50 551 1270 941 3267 1051 Ba 230.425 <5 <5 <5 <5 <5 <5 <5 <5 Be313.107 <5 <5 <5 <5 <5 <5 <5 <5 Ca 317.933 118499 135214 151205 144550159523 195419 136590 94084 Cd 214.440 10962 18099 5015 4911 23422 522426033 7266 Co 228.616 <5 <5 <5 <5 <5 <5 <5 <5 Cr 283.563 <10 <10 <10 <10<10 <10 <10 <10 Cu 324.752 422 3381 3266 4352 3283 3008 3942 8163 Fe238.204 1737 1442 663 1021 1042 868 1862 2484 K 766.490 5114 5447 71966302 6967 14557 23222 3932 Mg 279.077 364.0 313.3 239.6 360.26 509.15249.13 <50 1062.2 Mn 257.610 73.8 39.9 30.6 21.10 38.97 35.58 20.67142.37 Na 330.237 1059.01 1661.43 2214.97 2548.06 4577.18 2207.31 45261148.50 Ni 221.648 <20 <20 <20 <20 <20 <20 <20 <20 P 213.617 <100 <100<100 <100 <100 <100 <100 <100 Pb 220.353 934234 463786 450437 512057462415 327273 537730 595226 S 180.669 Sb 206.836 73.5 <50 <50 41.2 52.87<40 <70 138.6 Se 196.026 <50 <50 <50 <50 <50 <50 <50 <50 Si 251.611 362325 277 1017 2330 2107 5325 1865 TI 190.801 228.3 319.9 146.1 60.18299.21 80.4 382.1 153.4 V 292.402 <10 <10 <10 <10 <10 <10 <10 <10 Zn206.200 70098 54726 35044 15228 19536 24367 19956 42493 BC Stg 1 solids(uw) Run #10 Run #11 Run #12 Date 070799R 070799R 070799R I.D.R6&10BCS1C R11BCS1C R12BCS1C Sc 361.383 5.03 5.02 4.98 Ag 338.289 30.9429.1 13.7 Al 308.215 539.4 122.2 64.7 As 188.979 <10 <10 <10 B 249.772142.6 1154 469 Ba 230.425 43.7 <5 <5 Be 313.107 <5 <5 <5 Ca 317.933284250 162731 120377 Cd 214.440 17297 6579 6121 Co 228.616 <5 <5 <5 Cr283.563 <10 <10 <10 Cu 324.752 3200 4240 3795 Fe 238.204 2045 1284 1784K 766.490 8525 7254 5432 Mg 279.077 2094.64 312.42 243.31 Mn 257.61084.48 51.92 46.70 Na 330.237 1954.7 3102.8 1773 Ni 221.648 <20 <20 <20 P213.617 343.72 241.42 <100 Pb 220.353 462444 461901 448382 S 180.669 Sb206.836 87.01 52.8 69.4 Se 196.026 <50 <50 <50 Si 251.611 694 1866 955TI 190.801 191.6 146.1 127.2 V 292.402 <10 <10 <10 Zn 206.200 2554228751 24758

TABLE 10 BC Stage 2 filtrate Run #0 Run #1 Run #2 Run #3 Run #4 Run #5Run #6 Run #7 Date 070799r 070799r 070799r 070799r 070799r 070799r I.D.r0bcs2ff r1bcs2ff r2bcs2ff r3bcs2ff r5bcs2ff r7bcs2ff Sc 391.383 3.564.53 3.46 3.38 3.38 3.43 Ag 338.289 2.05 2.78 3.10 2.36 2.33 2.46 Al308.215 <5 <5 <5 <5 <5 <5 As 188.979 <1 <1 <1 <1 <1 <1 B 240.772 55.1060.54 62.58 62.24 64.46 68.98 Be 230.425 <1 <1 <1 <1 <1 1.12 Be 313.107<1 <1 <1 <1 <1 <1 Ca 317.933 142381 160572 152779 162257 157750 156478Cd 214.440 2.67 <1 <1 <1 <1 <1 Co 228.616 <1 <1 <1 <1 <1 <1 Cr 283.563<1 <1 <1 <1 <1 <1 Cu 324.752 <1 <1 <1 <1 <1 <1 Fe 238.204 <1 <1 <1 <1 <1<1 K 766.490 7408 8281 7591 8170 7703 8443 Mg 279.077 <5 5.73 5.22 11.397.50 <5 Mn 257.610 <1 1.71 <1 <1 1.30 1.39 Na 330.237 2107 2596 22811694 2573 2265 Ni 221.648 <10 <10 <10 <10 <10 <5 P 213.617 <5 12.29 <5<5 <5 <5 Pb 220.353 11.13 5.25 2.67 <1 <1 <1 S 180.669 Sb 206.836 <10<10 <10 <10 <10 <10 Se 196.026 <30 <30 <30 <30 <30 <30 Si 251.611 4.7515.62 4.99 7.87 10.51 6.56 TI 190.801 <15 <15 <15 <15 <15 <15 V 202.402<1 <1 <1 <1 <1 <1 Zn 206.200 9252 11498 9700 10996 9613 10655 1:10, K +Na + Ca + Zn 1:100 BC Stage 2 filtrate Run #8 Run #9 Run #10 Run #11 Run#12 Date 070799r 070799r 070799r 070799r 070799r I.D. r8bcs2ff r9bcs2ffr6&10bcs2ff r11bcs2ff r12bcs2ff Sc 391.383 3.43 3.40 3.45 3.44 3.48 Ag338.289 2.54 2.40 2.32 3.42 2.22 Al 308.215 <1 <1 <1 <1 0.01 As 188.979<1 <1 <1 <1 −2.32 B 240.772 68.31 67.35 64.29 64.09 66.67 Be 230.425 <1<1 <1 <1 0.99 Be 313.107 <1 <1 <1 <1 0.00 Ca 317.933 151799 159284155562 152877 138199 Cd 214.440 <1 <1 <1 <1 −0.04 Co 228.616 <1 <1 <1 <1−0.03 Cr 283.563 <1 <1 <1 <1 0.12 Cu 324.752 <1 <1 <1 <1 0.05 Fe 238.204<1 <1 <1 <1 0.03 K 766.490 7489 8306 7570 3373 7323 Mg 279.077 <5 <5 <5<5 0.97 Mn 257.610 <1 2.18 1.37 <1 0.55 Na 330.237 2018 1865 2856 38791491 Ni 221.648 <5 <5 <5 <5 0.01 P 213.617 <5 <5 <5 <5 <1.27 Pb 220.353<1 9.27 <1 <1 −0.13 S 180.669 Sb 206.836 <10 <10 <10 <10 0.42 Se 196.026<30 <30 <30 <30 1.00 Si 251.611 4.77 5.88 5.83 5.27 4.88 TI 190.801 <15<15 <15 <15 0.10 V 202.402 <1 <1 <1 <1 0.02 Zn 206.200 9456 10645 91789826 8963 1:10, K + Na + Ca + Zn 1:100

TABLE 11 American Metals Recovery Corp. ZnOx Lime ZnOx Lime Run #0&1 Run#2&3 Run #4&5 Run #7&8 Run #9&10&6 Run #11&12 Date before IS 070699R070699R I.D. r4&5znoxlime r7&8znoxlime Sc 361.383 4.68 4.76 4.87 Ag338.289 7.7 7.46 <1 Al 308.215 172.4 176.89 121.02 As 188.979 <1 <1 B249.772 <7 <5 <5 Ba 230.425 <1 <1 <1 Be 313.107 <1 <1 <1 Ca 317.933168438 160204 109598 Cd 214.440 <1 <1 <1 Co 228.616 <1 <1 <1 Cr 283.5631.4 1.09 <1 Cu 324.752 <1 <1 <1 Fe 238.204 150 133.19 91.82 K 766.490104.4 69.82 67.05 Mg 279.077 663.5 640.45 433.55 Mn 257.610 118 101.5971.65 Na 330.237 <5 <5 Ni 221.648 <5 <5 <5 P 213.617 — 17.66 13.07 Pb220.353 <15 <5 <5 S 180.669 1140.3 Sb 206.836 <10 <10 <10 Se 196.026 <30<30 <30 Si 251.611 828.3 666.41 455.81 TI 190.801 <15 <15 <15 V 292.402<1 <1 <1 Zn 206.200 23.3 81.43 39.72

TABLE 12 ZnOx Cake (uw) Run #0&1 Run #2&3 Run #4&5 Run #7&8 Run #6&9&10Run #11&12 Date 072199Y 072199Y 072199Y 072199Y 072199Y 072199Y I.D.R0&1ZnOXC 2&3ZnOXC R4&5ZnOXC R7&8ZnOXC R6&9&10ZnOXC R11&12ZnOXC Y371.029 5.05 5.17 5.07 5.07 5.06 5.04 Ag 338.289 <50 <50 <50 <50 <50 <50Al 308.215 233.7 181.8 169.2 143.6 135.8 223.8 As 188.979 <100 <100 <100<100 <100 <100 B 249.772 573.7 59.6 48.5 64.8 44.8 396.2 Ba 230.425 <15<15 <15 <15 <15 <15 Be 313.107 <1 <1 <1 <1 <1 <1 Ca 317.933 10968 2645020515 58116 37861 12393 Cd 214.440 <5 <5 <5 <5 <5 <5 Co 228.616 <15 <15<15 <15 <15 <15 Cr 283.563 <50 <50 <50 <50 <50 <50 Cu 324.752 <15 <15<15 <15 <15 <15 Fe 238.204 178.6 172.6 193.1 136.4 129.2 182.7 K 766.490517.5 1074.2 885.3 2582.6 1631.3 277.5 Mg 279.077 638.3 1084.6 926.3864.0 776.0 808.9 Mn 257.610 161.6 210.2 194.2 172.2 158.2 224.5 Na330.237 — — — — — — Ni 221.648 <15 <15 <15 <15 <15 <15 P 213.617 <50 <50<50 <50 <50 <50 Pb 220.353 330.4 139.8 <100 <100 <100 <100 S 180.669 <50378.6 <50 260.4 239.4 <100 Sb 206.836 <100 <100 <100 <100 <100 <100 Se196.026 <100 <100 <100 <100 <100 <100 Si 251.611 1883.0 1106.0 1002.0853.4 784.9 1739.6 TI 190.801 <50 <50 <50 <50 <50 <50 V 292.402 <15 <15<15 <15 <15 <15 Zn 206.200 519155 624049 574763 528840 463704 602939 ClZnO 99% 797057 797057 797057 797057 797057 797057

TABLE 13 American Metals Recovery Corp. ZnOx Filtrate ZnOx Filtrate Run#0&1 Run #2&3 Run #4&5 Run #7&8 Run #9&10&6 Run #11&12 Date 070699R070699R 070699R 070699R 070699R 081199 I.D. r0&1znoxff r2&3znoxffr4&5znoxff r7&8znoxff r9&10&6znoxff R11&12znoxff Sc 361.383 4.68 4.624.71 4.74 4.68 4.74 Ag 338.289 30.78 <1 <1 5.52 <1 <1 Al 308.215 <5 <1<5 <5 <5 <5 As 188.979 <1 <1 <1 <1 <1 <1 B 249.772 43.23 55.83 44.9742.83 46.50 50.82 Ba 230.425 <1 <1 <1 <1 <1 <1 Be 313.107 <1 <1 <1 <1 <1<1 Ca 317.933 135983 159312 138537 133089 141828 128107 Cd 214.440 1.193.69 1.66 <1 <1 <1 Co 228.616 <1 <1 <1 <1 <1 <1 Cr 283.563 <1 <1 <1 <1<1 <1 Cu 324.752 <1 <1 <1 <1 <1 <1 Fe 238.204 <1 <1 <1 <1 <1 <1 K766.490 6865 7965 6565 6578 6915 7106 Mg 279.077 <1 1.24 <1 <1 <1 <1 Mn257.610 <1 <1 <1 <1 <1 <1 Na 330.237 2314 870 2456 2058 2197 Ni 221.648<1 <1 <1 <5 <5 <5 P 213.617 <1 <1 <1 <5 <5 <5 Pb 220.353 6.07 <1 <1 <14.51 <2 S 180.669 Sb 206.836 <5 <5 <5 <5 <5 <5 Se 196.026 <5 <5 <5 <5 <59.45 Si 251.611 <3 3.99 <1 <3 <3 <3 TI 190.801 <15 <15 <15 <5 <5 <5 V292.402 <1 <1 <1 <1 <5 <5 Zn 206.200 713 9048 937 843 825 528 Cl

TABLE 14 Salt Cake Run #0&1 Run #2&3 Run #4&5 Run #7&8 Run #6&9&10 Run#11&12 I.D. r0&1nacl/kcl r2&3nacl/kcl r4&5nacl/kcl r7&8nacl/kclr6&9&10nacl/kcl Date 071399 071399 071399 071399 071399 Sc 361.383 Ag338.289 2.71 <2 <2 <2 <2 Al 308.215 29.08 6.07 10.99 11.90 17.57 As188.979 <15 <15 <15 <15 <15 B 249.772 19.26 15.91 31.43 41.95 54.72 Ba230.425 1.06 <1 <1 <1 <1 Be 313.107 <1 <1 <1 <1 <1 Ca 317.933 4681136807 66948 49472 164383 Cd 214.440 16.64 0.88 3.25 2.16 4.24 Co 228.616<1 <1 <1 <1 <1 Cr 283.563 7.96 <2 <2 <2 8.18 Cu 324.752 5.41 <1 <1 <12.06 Fe 238.204 861.16 47.51 49.58 18.55 199.01 K 766.490 2084 1724 32202537 2763 Mg 279.077 53.87 <15 <15 <15 13.55 Mn 257.610 57.70 3.71 2.911.51 16.64 Na 330.237 343720 313088 276826 422796 117539 Ni 221.648 1.18<1 <1 <1 5.73 P 213.617 <20 <20 <20 <20 <20 Pb 220.353 150.68 <5 5.61 <518.32 S 180.669 <15 <15 <15 <15 <15 Sb 208.836 <5 <5 <5 <5 <5 Se 196.026<10 9.46 10.10 <10 16.37 Si 251.611 41.72 22.30 15.30 15.56 54.60 TI190.801 <10 <10 <10 <10 <10 V 292.402 <2 <2 <2 <2 <2 Zn 206.200 1192200.30 523.29 522.07 568.94

TABLE 15 American Metals Recovery Corp. Zinc Oxide Cake Final FiltrateZnO filtrate Run #0&1 Run #2&3 Run #4&5 Run #7&8 Run #9&10&6 Date 062899063099 081999 081999 082499 ICP Time Stamp 2:27 I.D. r0&1znoff r2&3znoffr4&5znoff r7&8znoff r6&9&10znoff Sc 361.383 4.71 Ag 338.289 <2 <2 <2 <2<2 Al 308.215 <3 <3 <3 <3 <3 As 188.979 <15 <15 <15 <15 <15 B 249.772 <7<7 2.17 3.22 1.19 Ba 230.425 1.62 1.88 1.04 1.47 1.20 Be 313.107 <1 <1<1 <1 6633 Ca 317.933 4827 5929 6894 6933 0.28 Cd 214.440 <1 <1 0.14 <1<1 Co 228.616 <1 <1 <1 <1 <1 Cr 283.563 <2 <2 <2 <2 <2 Cu 324.752 <1 <1<1 <1 <1 Fe 238.204 <1 <1 <1 <1 0.19 K 786.490 23.89 76.92 165.04 63.7753.14 Mg 279.077 <1 87.75 61.24 95.97 121.74 MN 257.610 <1 1.91 2.01 <13.60 Ni 221.648 <1 <1 <1 <1 <1 P 213.617 <20 <20 <20 <20 <20 Pb 220.353<5 <5 <5 <5 3.09 S 180.669 30.19 20.11 28.86 63.50 Sb 206.836 <5 <5 <5<5 <5 Se 196.026 <10 <10 1.82 9.12 1.04 Si 251.611 <3 <3 <3 18.51 <3 TI190.801 <10 <10 <10 <10 <10 V 292.402 <2 <2 <2 <2 <2 Zn 206.200 68.5883.68 64.53 93.57 239.32 Cl

TABLE 16 American Metals Recovery ZnO Runs 7/1/99 ICP file 070199Analyte R2&3ZnOC R4&5ZnOC R0&1ZnOC Ag 73 BDL BDL Al 333 324 295 As BDLBDL BDL B BDL BDL 366 Ba BDL BDL BDL Be BDL BDL BDL Ca 3800 1240 10200cd BDL BDL BDL Cl 2000 2500 2500 Co BDL BDL BDL Cr BDL BDL BDL Cu BDLBDL BDL Fe 298 379 345 K BDL BDL BDL Mg BDL BDL 1310 Mn 302 * 275 323 Nanot available not available not available Ni BDL BDL BDL Pb BDL BDL BDLSb BDL BDL BDL Se BDL BDL BDL Si 1370 1990 1960 TI BDL BDL BDL V BDL BDLBDL Zn 786000 744000 765000 ZnO99% 784000 784000 777000 BDL = BelowDetection Limit

TABLE 17 KO-61 and coke addition only 12.3 Kg Carbon to 100 Kg of KO61Temperatures C. green ball 1000 1050 1100 1000 1050 1100 TTI solids TTI1 #1 TTI 1 #2 TTI 1 #3 TTI 1 #4 TTI 1 #5 TTI 1 #6 TTI 1 #7 Ag 338.289 <397.60 178.65 127.24 102.59 122.21 137.34 Al 308.215 2675 3917 4792 72455324 7180 10345 As 188.979 <15 <30 102.13 68.95 77.21 67.65 86.88 B249.772 736.29 1323 1529 2061 1361 1890 1986 Ba 230.425 233.17 360.62416.52 581.65 374.44 545.05 540 Be 313.107 <1 <1 <1 <1 <1 <1 <1 Ca317.933 40144 51832 67102 81495 55258 74269 77336 Cd 214.440 492.0627.00 37.00 68.97 29.97 40.36 45.68 Co 228.616 5.94 23.31 28.77 36.5823.28 31.65 34.99 Cr 283.583 783.36 1406 1699 2072 2088 2198 3530 Cu324.752 2062.48 3975 4825 6280 4255 5820 6008 Fe 238.204 82092 299906345600 472289 321522 434690 456668 K 766.490 12646 12727 15972 1302211369 11788 3203 Mg 279.077 5330 10847 12308 16254 11466 14893 16407 Mn257.610 8099 21102 24057 32125 22686 30159 32388 Ni 221.648 65.27 296.41366.19 446.72 308.09 393.58 452.73 P 213.817 990.05 1150 1301 1792 11271864 1826 Pb 220.353 15265 16609 15087 15223 15194 12253 8273 S 180.6699724 17979 16472 14235 14860 20756 19853 Sb 206.836 69.26 227.63 280.54337.66 233.61 315.93 349.74 Se 196.026 <10 <10 <10 <10 <10 <10 <10 Si251.611 467.12 1350.88 1429.11 2395.00 1112.76 1005.75 639.04 Ti 190.801<10 <10 <10 <10 <10 <10 <10 V 292.402 51.79 61.09 74.93 91.92 61.8682.62 95.79 Zn 206.200 154115 187804 131907 68801 173390 48516 46099 Cl31662 47032 45463 54059 53843 48749 3507 Na 29778 51037 46746 6323250827 61738 48003 Carbon 5.29-7.24 3.5-3.55 1.33-1.59 0.52 0.75-0.811.25-2.03 0.20-0.35 undissolved solids (mg/kg) 372247 135530 87636 9936486286 104974 66756 % undis solids 37.22 13.55 8.76 9.94 8.63 10.50 6.68Test Group 1 KO-61, Coke and Calcium Hydroxide form Leach cake 34 KgCa(OH)2; 12.3 Kg Carbon to 100 Kg of KO61 Temperatures green ball 10001050 1100 1000 1050 1100 TTI solids TTI 3 #1 TTI 3 #2 TTI 3 #3 TTI 3 #4TTI 3 #5 TTI 3 #6 TTI 3 #7 Ag 338.289 71.62 28.81 13.00 92.48 97.64113.61 54.36 Al 308.215 3712 7698 10317 11053 7914 9435 11338 As 188.979<30 <30 <30 <30 <30 <30 <30 B 249.772 915.27 1235 1663 1729 1439 16391584 Ba 230.425 211 333 443.40 471.40 381.01 427.40 416.94 Be 313.107 <1<1 <1 <1 <1 <1 <1 Ca 317.933 128859 186925 243870 255457 214479 241665237813 Cd 214.440 512.33 908.74 37.26 37.16 30.99 34.52 34.64 Co 228.61613.35 24.94 27.51 28.08 23.13 26.50 26.55 Cr 283.583 1707 2491 3419 36122840 3489 3290 Cu 324.752 2619 3672 4812 5018 4353 4912 4598 Fe 238.204193223 280025 364257 382183 322750 370845 345188 K 766.490 9249 89887885 2190 7334 3557 670 Mg 279.077 7421 10613 13953 14854 12035 1391413042 Mn 257.610 13590 19419 25603 26836 22849 26373 24385 Ni 221.648179.10 273.49 335.31 334.97 299.80 325.85 337.35 P 213.817 575 827 15021483 1322 1601 1311 Pb 220.353 14220 12594 7268 2056 7341 14030 2605 S180.669 9479 4935 6886 10662 13740 14797 16483 Sb 206.836 142.95 225.8881.10 259.50 235.74 268.46 238.60 Se 196.026 <10 <10 <10 <10 <10 <10 <10Si 251.611 752.27 405.74 350.64 453.06 437.34 425.62 827.80 Ti 190.801<10 <10 <10 <10 <10 <10 <10 V 292.402 42.56 69.01 103.01 88.14 69.9775.32 79.88 Zn 206.200 178049 99891 19899 2228 50962 25661 1010 Cl100743 32685 20023 4465 23191 10831 780 Na 26662 30373 30190 20961 2531427603 21676 Carbon 8.56-8.82 1.41-1.58 0.37-0.56 0.27-0.55 1.50-1.740.58-0.63 0.23-0.27 undissolved solids (mg/kg) 88123 47610 76667 4361140503 63488 49831 % undis solids 8.81 4.76 7.87 4.36 4.05 6.36 4.98sample description

TABLE 18 Test Group 2 KO61 plus Ca/Mg Reduction Tests to produce Fluxedsinter briquettes/pellets Preheat Time Heat Time Preheat Temp Heat Tempmg/kg mg/kg mg/kg mg/kg % Recovery-Normalized Test # minutes minutes °C. ° C. Zn Ca Cd Pb Zn Cd Pb KO61 221,000 114,000 495  17,900 4 + 4a 545 275 1030 3,260 228,000 32 162 98.8% 94.9% 99.3% 4b + 4c 5 65 275 1030614 214,000 30 109 99.8% 94.9% 99.5% 5 + 5a 5 15 275 1050 61,500 188,00024 15,800 73.1% 95.3% 14.7% 5b + 5c 5 25 275 1050 4420 254,000 38 453098.6% 94.5% 81.9% 5d + 5e 5 35 275 1050 1910 229,000 30 844 99.3% 95.2%96.3% 5f + 5g 5 45 275 1050 862 226,000 31 185 99.7% 95.0% 99.2% 5h + 5i5 55 275 1050 619 227,000 33 141 99.8% 94.7% 99.4% 6 + 6a 5 15 275 107027,100 338,000 66 17,200 93.4% 92.8% 48.4% 6b + 6c 5 25 275 1070 2580235,000 35 2700 99.1% 94.5% 88.3% 6d + 6e 5 35 275 1070 1890 234,000 336480 99.3% 94.8% 71.9% 6f + 6g 5 45 275 1070 925 231,397 102  3836 99.7%83.8% 83.2% 7 + 7a 5 15 275 1100 3230 245,549 29 5942 98.9% 95.7% 75.4%7b + 7c 5 20 275 1100 1860 229,655 31 6436 99.3% 95.0% 71.6% Round 2 MixVolumes Total Weight K061 Iron Ore Leach Cake CaO Coke Coat grams lbsgrams/test Test 4-7  100 18.065 15.723 10 1000 180.65 157.23 1337.882.94334 51.4569231 Test S1-S4 100 5.75 32.79  5 500 28.75 163.95 692.71.52394 49.4785714 Note: CaO from store bought calcined lime

1. A method of producing Simonkolleite comprising the steps of:providing a high concentration chloride solution including zinc chloridecomplex; and adding water to the high concentration chloride solution toreduce the chloride concentration in the high concentration chloridesolution so as to produce a reduced concentration chloride solutionhaving a specific gravity less than 1.45, wherein adding said waterresults in at least 30% of said zinc chloride complex precipitating outof the reduced concentration chloride solution as Simonkolleite and zincoxide and/or hydroxide.
 2. The method of producing Simonkolleite ofclaim 1 further comprising the step of: adding a base to the reducedconcentration chloride solution wherein adding said base results in atleast an additional 60% of said zinc chloride complex originally in saidhigh concentration chloride solution, precipitating out of the reducedconcentration chloride solution as Simonkolleite.
 3. The method ofproducing Simonkolleite of claim 2 wherein the method takes place atatmospheric pressure.
 4. The method of producing Simonkolleite of claim2 wherein the method takes place at a temperature less than 130 degreesCelsius.
 5. The method of producing Simonkolleite of claim 4 wherein themethod takes place at a temperature of less than 30 degrees Celsius andprecipitates approximately 95% of the zinc chloride complex out ofsolution.
 6. The method of producing Simonkolleite of claim 2 whereinthe reduced concentration chloride solution has a pH in a range of 3.5to 9.0.
 7. The method of producing Simonkolleite of claim 2 wherein thereduced concentration chloride solution has a pH in a range of 5.0 to7.5.
 8. The method of producing Simonkolleite of claim 3 wherein atleast 90% of said zinc chloride complex precipitates out of the reducedconcentration chloride solution as Simonkolleite.
 9. The method ofproducing Simonkolleite of claim 3 wherein said base comprises at leastone compound selected from the group consisting of calcium hydroxide,calcium oxide, sodium hydroxide, and potassium hydroxide.
 10. The methodof producing Simonkolleite of claim 2 wherein the reduced concentrationchloride solution contains zinc at a concentration greater than 500 ppm.11. The method of producing Simonkolleite of claim 2 wherein the reducedconcentration chloride solution contains zinc at a concentration greaterthen 10,000 ppm.
 12. The method of producing Simonkolleite of claim 3wherein the steps of adding the water and adding the base are conductedsimultaneously.
 13. The method of producing Simonkolleite of claim 2wherein the method process takes place at a temperature between thecrystallization temperature of the reduced concentration chloridesolution and the boiling temperature of the reduced concentrationchloride solution.
 14. The method of producing Simonkolleite of claim 2wherein the step of adding water to precipitate said zinc chloride isexothermic and the method further comprises the step of cooling thereduced concentration claim 2 chloride solution.
 15. The method ofproducing Simonkolleite of claim 2 wherein said Simonkolleite comprisesa portion of a filter cake and further comprising the step of washingthe filter cake with a hot solution of calcium chloride in water toremove excess potassium chloride from the filter cake.
 16. The method ofproducing Simonkolleite of claim 15 wherein the hot solution of calciumchloride in water is at a temperature between 70 and 130 degrees Celsiusand has a specific gravity between 1.41 and 1.45.
 17. The method ofproducing Simonkolleite of claim 2 wherein said Simonkolleite comprisesa portion of a filter cake and further comprising the step of washingthe filter cake with hot water to remove excess soluble chlorides. 18.The method of producing Simonkolleite of claim 17 wherein the hot wateris at a temperature between 20 and 100 degrees Celsius.
 19. A zinc oxidemanufacturing process comprising the steps of: providing a metalsbearing feed stock; providing a chloride leach solution with a specificgravity in a range of 1.45 to 1.55; reacting the metals bearing feedstock and the chloride leach solution to form a complex solutionincluding a metal chloride complex and calcium hydroxide; adding waterto the complex solution to reduce a chloride concentration in thecomplex solution so as to produce a cementation solution having aspecific gravity within a range of 1.40 to 1.49; adding zinc to saidcementation solution, wherein said zinc added to said cementationsolution is below the metal in the metal chloride complex on theelectrochemical replacement series such that the zinc added to thecementation solution will substitute with the metal in the metalchloride complex to form a zinc chloride complex and such that the metalsubstituted out of the metal chloride complex will cement out of thecementation solution; adding water to the zinc chloride complex toproduce a zinc chloride complex solution having a reduced chlorideconcentration with a specific gravity within a range of 1.37 to 1.45,wherein adding said water results in at least 30% of said zinc chloridecomplex precipitating out of the zinc chloride complex solution asSimonkolleite; adding a base to the zinc chloride complex solutionwherein adding said base results in at least an additional 60% of saidzinc chloride complex originally in said cementation solutionprecipitating out of the zinc chloride complex solution asSimonkolleite; adding water to the Simonkolleite to produce aSimonkolleite slurry; and adding a base to the Simonkolleite slurry toproduce zinc oxide having a purity of at least 98% and containing lessthan 1,000 ppm chlorides.
 20. The zinc oxide manufacturing process ofclaim 19 wherein the step of reacting the metals bearing feed stock andthe chloride leach solution takes place at atmospheric pressure, thechloride leach solution includes calcium chloride and a chloride ionconcentration greater than 10 molar, the temperature is above 65 degreesCelsius, and the pH of the complex solution is maintained above 3.5 andbelow 9.0.
 21. The zinc oxide manufacturing process of claim 19 wherein,in the step of reacting the metals bearing feed stock and the chlorideleach solution, the metal in the metal chloride complex comprises one ormore metals selected from the group consisting of zinc, lead, cadmium,silver, copper, tin, nickel, and a metal in the electrochemical seriesabove zinc.
 22. The zinc oxide manufacturing process of claim 19 whereinthe process utilizes water to maintain the specific gravity of thechloride leach solution in a range of 1.45 to 1.55.
 23. The zinc oxidemanufacturing process of claim 19 wherein the cementation solution has apH in a range of 5.0 to 8.0, and the step is conducted at atmosphericpressure and at a temperature greater than 65 degrees Celsius.
 24. Thezinc oxide manufacturing process of claim 19 wherein the steps of addingwater and a base to the zinc chloride complex solution takes place atatmospheric pressure and at a temperature of approximately 25 degreesCelsius, wherein the reduced chloride concentration solution has a pH inthe range of 3.5 to 9.0, and wherein said base comprises at least onecompound selected from the group consisting of: calcium hydroxide,calcium oxide, sodium hydroxide, and potassium hydroxide.
 25. The zincoxide manufacturing process of claim 19 wherein said Simonkolleitecomprises a portion of a filter cake and further comprising the step ofwashing the filter cake with a hot solution of calcium chloride in waterto remove excess potassium chloride from the filter cake.
 26. The zincoxide manufacturing process of claim 25 wherein the hot solution ofcalcium chloride in water is at a temperature between 65 and 130 degreesCelsius and has a specific gravity between 1.41 and 1.45.
 27. The zincoxide manufacturing process of claim 19 wherein the step of go adding abase to the zinc chloride complex solution takes place at a temperaturein a range of 140 to 200 degrees Celsius, in a pH range of 6.9 to 7.4,and at atmospheric pressure.